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By W. A. Gibson, Zay Jeffries

It has been known for a number of years that certain aluminum alloys could be hardened by quenching from a temperature of about 500' C. Immediately after quenching the total increase in hardness is not marked, but the hardness increases gradually so that after about 4 days, it has reached a maximum. The actual increase in hardness probably continues for weeks or even months, but in order to detect the differences after 4 days, it is necessary to make tests at rather long intervals. So far as is known to the authors, Alfred Wilm was the first to discover these properties in aluminum alloys.' The aluminum-copper alloys are amenable to this treatment and the addition of magnesium makes the heat-treatment effect more pronounced. Until recently heat treatment of aluminum alloys has been confined to the worked alloys in the form of sheet or extruded products. The alloy that has been most used in this connection is known commercially as duralumin and has about the following composition: copper not exceeding, 6 per cent.; magnesium not exceeding, 2 per cent.; manganese not exceeding, 1 per cent.; commercial aluminum,2 balance. A typically good alloy in this range consists of: copper, 4 per cent.; magnesium, 0.5 per cent.; manganese, 1 per cent.; commercial aluminum, balance. This alloy can be rolled into sheets or extruded; and in this condition it can be heated in the ordinary furnace atmosphere to a temperature near 500° C. and quenched in water. After about 4 days it will have the following physical properties: tensile strength 55,000 to 60,000 1b. per sq. in. (3866 to 4218 kg. per sq. cm.); elongation in 2 in., 15 to 20 per cent. No harmful effects of the furnace gases or even the air, if the heating is effected in an electric furnace in an air atmospherc.

Jan 1, 1921

By E. L. Tucker

IN THE metallurgy of precious metals, it has been standard practice for years to use cyanogen compounds, so it was but natural that early investigators in the field of flotation should consider these compounds as possible flotation reagents. In Australian Patent 9508, of 1913, alkali cyanides were described as suitable reagents for the separation of heavy sulfide minerals, the reagents being employed either in an alka-line or neutral circuit. In U. S. Patent 1236501, of 1917, to John M. Tibbet, the use of cyanide in conjunction with soda ash or lime is described. One of the claims of this patent reads, "An improvement in the cyanide process for the treatment of ores comprising, mixing the ore with cyanide solution and oil and alkali: emulsifying the oil in the presence of the ore, the alkali and the cyanide solution, floating a portion of the values in the pulp by producing gas bubbles in the mixture-and removing the values floated." More recently, Sheridan and Griswold, in U. S. Patents 1421585 and 1427235, describe a somewhat similar process and have, successfully introduced it in. the mill of the Timber Butte Milling Co. U. S. Patent 1429544, of 1922, to Stevens, describes the use of double cyanides of copper, zinc, and lead. The success of cyanide as a reagent in the differential flotation of lead and zinc sulfides in highly complex ores is proof that these patents have a solid foundation.. Although the technique of their application to particular ores has been worked out in many cases, no one, so far as is known, has attempted to describe the nature of the action of the cyanide or to determine the ideal conditions for its, application. It was the purpose of the present investigation to gather information on these points. Naturally this would have been an exceedingly ambitious.

Jan 8, 1925

By Frederick O?Neil

SCOPE OF THIS REPORT THE object of this paper is to describe the present status and possibilities of liquid oxygen as an explosive based upon the investigations, research and practical work of the Ingersoll-Rand Co., and of the Air Reduction Co., from early in 1922 to date. The following work was done. 1. A study of the history and physical theory of L. 0. X. as an explosive. (No attempt will be made in this paper to give the history or theory, as this has already been covered by the literature on the subject.) 2. An investigation of the use and the practice with L. 0. X. in the iron mines of Lorraine where L. 0. X. has its largest application in Europe. Although L. 0. X. is extensively employed in the; iron mines of Lorraine, many of the mines using it exclusively, and blasting with it a total of approximately three million tons of iron ore per year, the fact that the mining practice in Lorraine1 is to drill comparatively shallow holes and blast only a few holes at a time, renders the data obtained there of very little value in estimating the possibility of replacing dynamite with L. 0. X. in American mining methods. In fact, in the iron mines in Lorraine, L. 0. X. replaced black powder, or its equivalent. 3. Laboratory research work on properties of absorbent materials. 4. Work in the Bureau of Mines explosive laboratory at Pittsburgh, Pa., on the properties of L. 0. X. cartridges. 5. Practical work underground with L. 0. X. cartridges in the Witherbee-Sherman iron mines at Mineville, N. Y. 6. Practical work in the quarries of the Calcite Quarry Co., Myers-town, Pa. 7. Practical blasting in several quarries and open-pit mines.

Jan 2, 1926

By H. F. Silliman, Daniel R. Hull, Earl W. Palmer

THE brass-rolling industry has not had a great deal of experience with antimony in its product. There have been some recent excursions with antimony as a corrosion inhibitor in tubes, but in sheet brass it has been regarded with apprehension and generally avoided. This is not to say that [ ] contaminations have never occurred, and probably a great many brass mill men have had their own ideas on the effect of such contamination. Our own views were based upon certain experiences some years ago with strip castings that split lengthwise in the early stages of rolling. The nature of this splitting is shown in Fig. I, from a photograph taken in 1922 as a record of 20 bars of 65-35 common brass made with an antimonial brand of copper. These bars were cast 6 ¾ in. wide by 1 in. thick by 45 in. long. In the first pass through the rolls, 19 of the 20 split lengthwise throughout their length, while each half curved outward in a wide arc as it emerged from the rolls. [ ] In those days, if splitting was not a frequent occurrence, it was not unfamiliar. One can recall periods of short duration when large amounts of common brass would behave in this way. There was nothing imaginary about these spells. The affected metal would split lengthwise in the breaking rolls, and no maneuvering of mill practice made any difference. The condition would descend like a plague, demoralize the mill for several days, and disappear. It is doubtful if anyone knew the cause-however many explanations may have been current. In 1921 a quantity of split brass was traced to the use of fired small-arms

Jan 1, 1943

By M. M. Austin

Pure metallic tantalum, from a practical standpoint, is one of the newer developments in the metal field. Although it was used as filament in incandescent lamps in 1906, only within the last, five years has the metal been available in sufficient quantity and suitable form for more general application. In addition to corrosion-resisting properties and high melting point, the metal has an ability to withstand unusual amounts of cold working. It affords some striking demonstrations of block movement during plastic flow. Strength and Ductility of Tantalum Since there is no good way of defining working properties numerically, illustrations will be used. It is possible to roll a bar of pure tantalum 0.4 in. thick to sheet 0.001 in. thick, without heating or annealing. The sheet produced will be sound and may be bent upon itself at right angles to the direction of rolling. Longitudinally the sheet will be somewhat brittle. A disk of annealed metal may be cupped and drawn to the form of a capsule 3/16 in. dia. by 146 1? in. long. If carefully handled this capsule can be further deformed by drawing to a tube having an outside diameter of 0.050 in. Disks of tantalum may be spun into tubes of 39 in. dia. and 1½ in. long. In the form of rod, reductions may be made from 0.4 in. to 0.001 dia. In the latter process, it is necessary to heat the wire four or five times to 500" C., at intervals between 0.040 and 0.001 in. Although the melting point of tantalum is 2850" C., these heatings at 500" C. are sufficient to distinctly anneal it. Published figures on the strength and ductility of tantalum are not particularly accurate. This is due to the fact that these properties vary widely, depending on the extent to which the metal is contaminated by gas. There are no sharp critical points in these changes with gas content; samples of metal will vary from a springy or actually brittle condition to one of leadlike softness. Samples that have had a prolonged heating at high temperatures in a good vacuum, under conditions such as would be provided in vacuum-tube experimentation, will yield a product of about the hardness of annealed copper. The usual commercial metal will have the hardness of annealed mild steel.

Jan 1, 1930

By Graham John Mitchell

In May, 1919, a crosscut on the 1400-ft. level of the Briggs mine, a Calumet, & Arizona property at Bisbee, Ariz., penetrated a deposit of pyrite and chaleopyrite that had replaced quartzite and limestone and was cut, especially in the quartzite, by numerous fractures. As the drift advanced through these fractured portions, they were found to be free from copper-sulfate minerals although '(sooty" chalcocite was present in the fractures and also as coatings and replacements of pyritc. Howcver, sulfate soon made its appearance in certain fractures, filled them, and began to form stalactites. Several centers of incipient growth were noted for subsequent study, observations in the drift at frequent intervals during the past 17 months affording the opportunity of determining the rate of formation of the stalactites. The one to attain the greatest growth during the period of observation measured 27 in. long and averaged 1 in. in diameter. A close study showed it to be composed of small interlocked crystals of chalcanthite radiating from a small central tube-like opening. The trielinic form of the crystals was evident with prism and pinacoid faces recognixable. Assuming that the supply of copper-sulfate solution was constant (luring the 17 months—-and such an assumption is supported by the fact that the number of drops of excess solution falling, per minute, from the tip of the stalactite at the times of observation checked closely—thc stalactite, it has been estimated, grew at the rate of 1.58+ in. per month. DISCUSSION James I?. Kemp," New Yolk City.—J. S. Curtiss, in the monogr:tpll on Eureka, Ncv., say:: that a stalactite formed in limestonc grew 5/8 in. in three weeks and continued to grow as long as it had a drop of water or solution at the lower end. Another grew at the rate of 3/8 in. in the same time. In the case cited by Mr. Mitchell, the growth was about 136 in. in three weeks. A year ago we had an instance of the formation of cubes of galena on a railroad spike. A mine in northeastern Oklahoma had been idle about two years but on account of pumping of neighboring properties, the circulation of water was active throughout the drift. A cube of galena,

Jan 1, 1922

By B. E. Lindsly

The withdrawal of gas and/or oil in appreciable quantity from a natural oil reservoir causes the pressure within the reservoir to diminish, and if the oil is completely saturated with gas, this dissolved gas begins to evolve from solution. The liberation of gas from solution in oil causes the oil to change in several ways, the principal changes being: (1) reduction in volume or "shrinkage"; (2) increase in specific gravity (decrease in degrees A.P.I.); (3) decrease of energy in the oil; (4) increased viscosity; (5) increased surface tension. As the writer's investigations have been concerned more with the first three items than with changes in viscosity and surface tension the following discussion relates particularly to the volumetric shrinkage of the oil, to the increased specific gravity, and to the decreased energy in the oil, each caused by the liberalion of dissolved gas. One of the most surprising facts that have developed in the study of high-pressure well-head and "bottom-hole" samples is the great reduction in volume that occurs in some oils when the naturally dissolved gas is liberated. High-pressure samples taken from well heads in the Oklahoma City, Kettleman Hills and Ventura, Calif., fields decreased in volume from 11.2 to 40.5 per cent when the dissolved gas was liberated in reducing the pressure on the sample to atmospheric. The amount of shrinkage an oil undergoes when gas is liberated depends upon the quantity of gas liberated and also upon its composition. A given volume of the heavier gases, propane and butane for example, will occupy more space in the oil than an equal volume of methane. The reason for this probably is that the molecules of the heavier gases are larger than those of the lighter gases. When gases are not in solution, and particularly when they are at low pressures, the distance between the gas molecules is so much greater than the dimensions of the molecules themselves that the mere size of the molecules has little effect upon the space occupied by the gas, Avogadro's hypothesis, which in effect states that all gases under the same conditions have the same number of molecules per unit volume, relates to this phenomenon. The fact that the heavier gases occupy more space than the lighter gases in the oil, and

Jan 1, 1934

C.A. MITKE, *Bisbee, Ariz. (written-discussion-? ).--The advantages of the incline-slicing system have' been ably pointed out by Mr. Scott in his paper in the November, 1917, Bulletin, and also by Mr. P. B. Scotland in the Engineering & Mining Journal, Apr. 7, 1917. From the standpoint of efficiency and economy this method is far superior to the top-slice system as generally practised. However, as regards ventilation, and the danger of mine fires, it is just as necessary to use a mechanical ventilating system in mines worked by this method as it is in those worked by the horizontal top-slice system. Mechanical ventilation not only improves working conditions in top-slice stopes but, if properly designed for the particular mine in question, permits an immediate attack to he made in the event of a mine fire with a greater degree of safety to the men. In the course of time, incline top-slicing will probably take the place of the horizontal top-slicing system just as the incline cut-and-fill has superseded the horizontal cut-and-fill stoping method. While incline top-slicing has been tried in several districts on a very limited scale with inconclusive results, the Coronado is the-only large mine at which it has yet been thoroughly tested, and there it proved to be a complete success. As Mr. Scott's figures show, the tonnage has been more than doubled, with a simultaneous reduction in costs. As has been the history with all other innovations in mining, considerable hesitation will probably be shown in introducing this method in other mines. For example, when the incline cut-and-fill system was put forward as a substitute for the horizontal cut-and-fill stoping method many mining men of good repute were doubtful of its success and post-poned its introduction as long as possible, giving as their reason certain geologic conditions which would make the incline cut-and-fill impracticable. At the present time, most of the mining companies in the South-west, who in the past had used the horizontal cut-and-fill system, have already changed over, or are in the process of changing, and have proved that in the majority of cases horizontal cut-and-fill stopes can be successfully operated under the incline cut-and-fill system. Large horizontal cut-and-fill stopes can now be found in only a few exceptional cases, and

Jan 3, 1918

Consolidated Gold Fields Limited published annually an excellent reference, GOLD 1978, GOLD 1979, giving the Supply and Demand for Gold, Recent Developments and Prospects, Soviet Gold Production, etc. This outstanding reference is authored by Christopher Glynn, Investment Department, London, Great Britain. The 1978 Report marked the tenth anniversary of Gold Field's review of world gold market developments; and from it and the 1979 report, the following information was derived. The recovery of gold is widespread in the Soviet Union. Originally, gold was mined principally in the Urals, but in the last few years several major deposits have been discovered in the southwestern areas of the Soviet Union. One of the most dramatic discoveries has been the Muruntau deposit in Western Uzbekestan. This deposit, discovered in 1958, is reported to be the largest and among the lowest cost producers in the Soviet Union. It is located in Armenia, approximately 750 km (465 miles) northeast of the Iran-Afghanistan border, lying in the center of the Kyzylkurn (Red Sands) desert at the foot of Muruntau Mountain. Free gold occurs in both fine and coarse size in quartz veins up to 20 m (66 ft) thick associated with a small percentage of sulfides, mainly arsenopyrite. The ore zone constitutes a large gold-enriched stock work of about 4 sq krn (1.5 sq miles). The development of Muruntau was a major undertaking with open -pit development and construction of the mill and refinery complex started early in 1967, and the first gold bar poured in mid-1969. At the end of the first construction stage in 1970, the mill building was an imposing structure about 38 m (125 ft) high and 330 (1,080 ft) in length housing, probably, 11 parallel milling circuits. The first construction stage also included six 50 m (165 ft) diameter thickeners and 30 Pachuca tanks. The original mill complex was tripled in size after large additional gold reserves were discovered at a depth of 600 m (1,970 ft); and the deposit is now being exploited by both open- pit and underground workings. Ore is ground in semi-autogenous mills with the free gold being recovered into a high-grade concentrate on shaking tables. Table tailings are purnped to duplex spiral classifiers in closed circuit with the grinding mills. Classifier overflow is then thickened before being transferred to the Pachuca building where cyanidation takes place. An ion exchange resin is then added in a countercurrent

Jan 1, 1981

By G. Sandoz, C. A. Siebert

OIL quenching of steel, in good commercial practice, is carried out using a great deal of agitation in the quenching bath. Many of the tests reported in the literature represent the results obtained on steels quenched into still oil. Scott1 describes three stages of cooling in liquid quenching baths, namely: "State `A'-Vapor Blanket Cooling-In this, the first stage, a thin stable vapor film surrounds the hot metal. Cooling is by conduction and radiation through the gaseous film and is therefore relatively slow. "Stage `B'-Vapor Transport Cooling- Termination of the `A' stage is marked by wetting of the metal surface. Vapor forms copiously in bubbles and is carried away by gravity and convection currents to condense in cooler surrounding liquid. This is the fastest stage of cooling. "Stage `C'-Liquid Cooling--The `B' stage ends as the surface temperature of the metal approaches the boiling point of the quenching liquid. Vapor no longer forms, so cooling is by conduction and convection and the temperature difference is greatly decreased. This mode of cooling is slower than that of the `B' stage." Mechanical agitation of the quenching medium will accelerate all three stages of cooling. The greatest effect, however, will be noted on stage `B' because the removal of the bubbles from the surface of the steel will not be entirely dependent upon gravity and natural convection currents. The method employed in this investigation for evaluating the oils was the end quench test. The various oils were substituted for water as the quenching liquid and hardness determinations were made at various distances from the quenched end. Most of the tests were run using a 3/8-in. orifice with a 2 ½ -in. stream height. However, to check the effect of other velocities two additional conditions were used, namely: (I) A 3/8-in. orifice and 1-in. stream height, and (2) a ½ -in. orifice and a 3 ½ -in. stream height. In all cases the samples were suspended so that the quenched end was ½ -in. above the orifice. The steel used was an NE 9450 having the following analysis: [ ] Some tests were run on an NE 8740, 3312, and a 4340 steel. The results obtained on these steels confirmed those obtained on the NE 9450 and have therefore been omitted' from this paper. The test bars were machined from 1 1/8 -in. stock to 1-in. diam by 4 in. long.

Jan 1, 1948

By Charles Wurms, Robert Steinitz

DURING an investigation to determine the compatibility of several metals with tungsten, the degree of reaction between tungsten and lead was determined. A literature survey showed considerable disagreement between the results of various earlier investigators. Information presented by M. Hansen1 states that, according to S. Inouye,2 tungsten has very poor resistance against attack by molten lead, even at the low temperature near the melting point of lead. A phase diagram in the same reference indicates a solubility of 30 wt pct W in molten lead at 1300°C. Contrary evidence was shown by Parkman and shepard3 in their investigation of materials for use in heat transfer systems containing liquid lead alloys; they performed experiments at 980°C for times up to 50 hr without noticing any appreciable attack on the tungsten. Tungsten sheet was immersed in molten lead contained in Vycor tubes evacuated to at least 10-4 mm Hg. R. Bernard4 and R. I. Jaffee5 also indicate that no reaction occurs between molten lead and tungsten. However, no experimental data were included in these reports. In order to settle the question of the solubility of tungsten in molten lead, tungsten powder with an average Fisher subsieve particle size of 6µ was pressed in a double-acting die at 6 tsi to small compacts. They were then sintered in an atmosphere of hydrogen for 2 hr at 1700°C, resulting in a sintered density of 64 pct of theoretical with final dimensions of 0.260 in. in height and 0.245 in. in diameter. For all tests, the tungsten pellets were immersed in molten lead. Table I is a summary of the experimental procedures showing atmospheres, temperatures, and times. The first experiments were performed at 650°C; at this temperature, the liquid lead should wet the porous tungsten better than at its melting point, and penetrate into the pores. The lead was heated in a graphite crucible with a graphite cover. No protective atmosphere was used, as it was expected that the formation of CO would be sufficient to prevent oxidation. This, however, proved to be wrong, and all tungsten samples showed a slight oxide film after treatment. The tungsten samples were immersed in the molten lead for times of 5 min, 10 min, 1/2 hr, 1 hr, and 2 hr. After this time, the melt was poured into a form and the

Jan 1, 1962

By C. A. Siebert, G. Sandoz

Oil quenching of steel, in good commercial practice, is carried out using a great deal of agitation in the quenching bath. Many of the tests reported in the literature represent the results obtained on steels quenched into still oil. Scott1 describes three stages of cooling in liquid quenching baths, namely: "State 'A'— Vapor Blanket Cooling—In this, the first stage, a thin stable vapor film surrounds the hot metal. Cooling is by conduction and radiation through the gaseous film and is therefore relatively slow. "Stage 'B'—Vapor Transport Cooling— Termination of the 'A' stage is marked by wetting of the metal surface. Vapor forms copiously in bubbles and is carried away by gravity and convection currents to condense in cooler surrounding liquid. This is the fastest stage of cooling. "Stage 'C'—Liquid Cooling—The 'B' stage ends as the surface temperature of the metal approaches the boiling point of the quenching liquid. Vapor no longer forms, so cooling is by conduction and convection and the temperature difference is greatly decreased. This mode of cooling is slower than that of the 'B' stage." Mechanical agitation of the quenching medium will accelerate all three stages of cooling. The greatest effect, however, will be noted on stage ' B' because the removal of the bubbles from the surface of the steel will not be entirely dependent upon gravity and natural convection currents. The method employed in this investigation for evaluating the oils was the end quench test. The various oils were substituted for water as the quenching liquid and hardness determinations were made at various distances from the quenched end. Most of the tests were run using a 3/8-in. orifice with a 21/2-in. stream height. However, to check the effect of other velocities two additional conditions were used, namely: (I) A 3/8-in. orifice and I-in. stream height, and (2) a 1/2-in. orifice and a 31/2-in. stream height. In all cases the samples were suspended so that the quenched end was 1/2-in. above the orifice. Some tests were run on an NE 8740, .3,312, and a 4340 steel. The results obtained on these steels confirmed those obtained on the NE 9450 and have therefore been Omitted from this paper. The test bars were machined from 1/8-in. stock to I-in. diam by 4 in. long.

Jan 1, 1949

By G. Sandoz, C. A. Siebert

Oil quenching of steel, in good commercial practice, is carried out using a great deal of agitation in the quenching bath. Many of the tests reported in the literature represent the results obtained on steels quenched into still oil. Scott1 describes three stages of cooling in liquid quenching baths, namely: "State 'A'— Vapor Blanket Cooling—In this, the first stage, a thin stable vapor film surrounds the hot metal. Cooling is by conduction and radiation through the gaseous film and is therefore relatively slow. "Stage 'B'—Vapor Transport Cooling— Termination of the 'A' stage is marked by wetting of the metal surface. Vapor forms copiously in bubbles and is carried away by gravity and convection currents to condense in cooler surrounding liquid. This is the fastest stage of cooling. "Stage 'C'—Liquid Cooling—The 'B' stage ends as the surface temperature of the metal approaches the boiling point of the quenching liquid. Vapor no longer forms, so cooling is by conduction and convection and the temperature difference is greatly decreased. This mode of cooling is slower than that of the 'B' stage." Mechanical agitation of the quenching medium will accelerate all three stages of cooling. The greatest effect, however, will be noted on stage ' B' because the removal of the bubbles from the surface of the steel will not be entirely dependent upon gravity and natural convection currents. The method employed in this investigation for evaluating the oils was the end quench test. The various oils were substituted for water as the quenching liquid and hardness determinations were made at various distances from the quenched end. Most of the tests were run using a 3/8-in. orifice with a 21/2-in. stream height. However, to check the effect of other velocities two additional conditions were used, namely: (I) A 3/8-in. orifice and I-in. stream height, and (2) a 1/2-in. orifice and a 31/2-in. stream height. In all cases the samples were suspended so that the quenched end was 1/2-in. above the orifice. Some tests were run on an NE 8740, .3,312, and a 4340 steel. The results obtained on these steels confirmed those obtained on the NE 9450 and have therefore been Omitted from this paper. The test bars were machined from 1/8-in. stock to I-in. diam by 4 in. long.

Jan 1, 1949

By E. O. Bennett

When oil is takcn from a subsurface structure it is generally accompanied by gas. The gas thus produced represents the 1ighter hydrocarbons present in the original petroleurn accurnulation, which arc, insoluble at the reduced pressures created around a well when it is producing. In some cases the gas accompanying the oil may come from a gas-bearing stratum not in immediate contact with the oil. The amount of gas from the formation accompanying each barrel of oil is called the natural gas factor. Energy Causing Well to Flow The force causing the flow of fluid into a well when it is drilled into an oil-bearing structure is the pressure differential between the well bore and the formation. A pressure differential between the formation and the well may be set up in three different ways, as follows: (1) pressure due to gas in solution or in contact with the oil, (2) hydrostatic pressure due to water in contact with the oil (it may be either bottom or edge water); (3) pressure difference due to forcc of gravity. The greatest source of energy for producing oil from most oil fields, and the one over which the greatest control can be exerted, is the poten-tial energy in the gas under pressure in the formation. In any pool there is but a definite amount of gas available. Part of this gas may be in solution and a part in contact with oil. The amount of gas in solution in each bard of oil depends on the pressure and temperature in the formation and the characteristics of the oil. Whether the gas is in solution or in contact with the oil it is necessary to use the energy it, contains as efficiently as possible in order to obtain the greatest ultimate recovery from the pool. Bush1 states: Gas being the most efficient aid in moving the oil, it is important to prevent more of it from escaping than is necessary to remove the maximum amount of oil from the sands. The conservation of this energy is far more important to the producer of oil than his returns from sales of gas to distributing companies. The flowing life of the well is the most profitable period, and keeping the gas-oil ratio low, or keeping the

Jan 1, 1928

By T. Sterry Hunt

The deposit of rock-salt which is known to exist along the eastern shore of Lake Huron, in the province of Ontario, has lately been more completely explored than before, by a boring with a diamond drill, put down by Henry Attrill, Esq., of New York City, and the results obtained are so important in every way that I make no apology for presenting them to the Institute of Mining Engineers. I may be permitted to preface an account of this remarkable exploration, and of its results, with a historical sketch of the discovery and development of this salt-region. It was in December, 1865, that a boring was begun near the town of Goderich, in the hope of finding petroleum. In this the adventurers were disappointed, but, after passing through about 800 feet of limestone, they encountered a series of variegated marls, in which, at a, depth of 964 feet from the surface, a bed of rock-salt, 30 feet in thickness, was met with in May, 1866. The boring was carried to a depth of 1010 feet, ending in hard rock, and yielded, by pumping, a very pure saturated brine when examined by me in August of the same year. In the report of the Geological Survey of Canada for 1863-66, published early in 1867, I described this salt-well, with many geological details, and gave an analysis of the brine. In the next three years a considerable number of wells was sunk in and around Goderich, and numerous trials were made in various other parts of the region. Salt was found at Kincardine, thirty miles north-northeast from Goderich, at a depth of about 900 feet, and also at Clinton, thirteen miles southeast, at 1180 feet. Records of these wells, with analyses of the brine from them, were given by myself in a subsequent report of the Geological Survey, 1866-69, published in 1870 (pp. 211-244), together with accounts of various unsuccessful borings in the neighborhood, analyses of the brines from the various wells (including one analysis by Dr. Goessmann), with many details of the salt-manufacture at Goderich, at Syracuse, New York, and at Saginaw, Michigan. The geological character of the region was there discussed at length, and it was shown that the salt here occurs in the Onondaga or Salina formation, which

By T. Sterry Hunt

THE deposit of rock-salt which is known to exist along the eastern shore of Lake Huron, in the province of Ontario, has lately been more completely explored than before, by a boring with a diamond drill, put down by Henry Attrill, Esq., of New York City, and the results obtained are so important in every way that I make no apology for presenting them to the Institute of Mining Engineers. I may be permitted to preface an account of this remarkable exploration, and of its results, with a historical sketch of the discovery and development of this salt-region. It was in December, 1865, that a boring was begun near the town of Goderich, in the hope of finding petroleum. In this the adventurers were disappointed, but, after passing through about 800 feet of limestone, they encountered a series of variegated marls, in which, at a, depth of 964 feet from the surface, a bed of rock-salt, 30 feet in thickness, was met with in May, 1866. The boring was carried to a depth of 1010 feet, ending in hard rock, and yielded, by pumping, a very pure saturated brine when examined by me in August of the same year. In the report of the Geological Survey of Canada for 1863-66, published early in 1867, I described this salt-well, with many geological details, and gave an analysis of the brine. In the next three years a considerable number of wells was sunk in and around Goderich, and numerous trials were made in various other parts of the region. Salt was found at Kincardine, thirty miles north-northeast from Goderich, at a depth of about 900 feet, and also at Clinton, thirteen miles southeast, at 1180 feet. Records of these wells, with analyses of the brine from them, were given by myself in a subsequent report of the Geological Survey, 1866-69, published in 1870 (pp. 211-244), together with accounts of various unsuccessful borings in the neighborhood, analyses of the brines from the various wells (including one analysis by Dr. Goessmann), with many details of the salt-manufacture at Goderich, at Syracuse, New York, and at Saginaw, Michigan. The geological character of the region was there discussed at length, and it was shown that the salt here occurs in the Onondaga or Salina formation, which

Jan 1, 1877

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By T. L. Joseph

Blast furnaces are most efficient thermally when the CO2 in the top gas is highest. Oxygen introduced in the air blast is converted to CO in the combustion zones. The extent to which CO, generated in front of the tuyeres, is converted later to CO2 depends upon the ratio of ore to coke in the charge and the degree to which CO2, formed by reduction of iron oxides with CO, reverts to CO by the secondary reaction C + CO2 = 2CO. About 300 to 400 lb. of carbon' per ton of pig iron is normally involved in the solution-loss reaction or direct reduction of FeO with carbon, which produces the same over-all results; that is, lower CO2 in the top gas. Rich2 has recently reported a reduction in fuel consumption at slower rates of blowing on two furnaces in spite of higher heat losses per ton of iron. Lower fuel consumption at slow operating rates was due to more complete reduction of the ore that resulted in higher CO2 in the top gas and consequently more efficient use of carbon. Kinney3 has reported substantial savings in fuel as a result of crushing and sizing the ore charge. Improvements in practice brought about by sintering magnetite concentrates, flue dust and fine ores have been due to the establishment of conditions favoring more complete reduction of the ore in the upper part of the furnace. The extent to which oxides and fluctuations in composition in pig iron affect the quality of steel and finished product needs further study. There is evidence that more thorough preheating and reduction of the ore in the upper part of the furnace tend to give a better quality of iron. Bogatzkii4 found from laboratory work that the amount of oxygen entering electrolytic iron, melted under various slags, was directly proportional to the concentration of free FeO in the slag and inversely proportional to the basicity of the slag. Herty and Gaines5 concluded that the silicate content of open-hearth steel before deoxidation is related to the concentration of silica in the pig-iron charge, indicating that most of the persistent inclusions are silica from the iron. They also found that

Jan 1, 1936

By T. C. Campbell

Copper is extracted from the Chuquicamata oxide ore by a hydro-metallurgical process. The ore is crushed to 3/8-in. sizing, and leached with a sulfuric acid electrolyte. Chlorine is precipitated. and the ferric iron reduced in the enriched electrolyte, after which the copper is recovered by electrolysis with insoluble Chilex and lead-antimony anodes, the spent electrolyte being returned to leaching. Cathodes are melted and refined in market furnaces and cast into commercial wire-bar and cake shapes. Sulfuric acid for the process is supplied by the brochantite in the ore. Water for washing the ore is advanced through the solution system and finally after cutting by electrolysis to from 6 to 16.0 grams per liter is completely stripped by the cuprous chloride method and run to waste. The cuprous chloride so obtained plus that resulting from precipitation of chlorine from strong solution is dissolved in ferrous chloride brine and the copper cemented on scrap iron. The cement so obtained is in part used to reduce ferric iron in electrolyte, and as it is of exceptional purity is also furnace refined and cast into an exceptional quality of fire-refined copper. Fig. 1 is flow sheet of the operation. The reduction plant is composed of seven divisions as follows: (1) crushing, (2) leaching, (3) tailings disposal, (4) dechloridizing, (5) sulfur dioxide treatment, (6) electrolytic tank house, (7) smelting and melting. A brief description of each of these divisions will be given. Crushing Ore that is frequently blocky and up to 5 ft. in section is crushed in four stages to 3/8-in. mesh. The first stage of crushing is called primary and the next three stages are known as secondary crushing. The first stage is accomplished with two 60-in. Superior McCully gyratory breakers by which the ore is broken to 9 in. The second coarse gyratory breakers crush the ore to 3-in. sizing. There are seven No. 10 McCully gyratories in this service, for one of which recently has been substituted a No. 7 cone crusher. The third and fourth stages of crushing are accomplished by fifty 48-in. disk crushers. In the third stage 14 disk crushers break the ore to 1 in., and in the last stage 36 disk crushers turn out the final product of 0.371-in. sizing. Reference to the flow sheets of Figs. 2 and 3

Jan 1, 1934

By E. R. Johnson

The copper concentrator of the Braden Copper Co. is at Sewell, Chile, on the westem flank of the main Cordillera of the Andes, at an air distance of approximately 50 miles southeast of Santiago, the capital of Chile. The region is mountainous and space is at a premium because of the rough terrain and the danger of snowslides. The ore, a porphyry copper, is mined by a block-caving method in the El Teniente mine. It is hauled to the crushing plant, a distance of 2.6 miles, through a tunnel because of the severe climatic conditions. The usual large, coarsely set, primary reduction machines found in many ore-treating plants are not required because of the mining method employed and the limiting 12-in. spacing of the grizzlies at the head of each ore pass in the mine. It is the purpose of this paper to describe the numerous improvements made in the crushing section of this large flotation concentrator during the past I5 years to increase the efficiency of the plant and to raise the daily output from around 20,000 to 30,000 tons. The reader should keep in mind that all changes were made within the existing building because of the limited space and the high cost of the heavy mill construction required to withstand the snow loads. The present plant, as well as the former, operates on a 24-hr, 7 days a week basis. The crushing plant commenced operations at its present site in 1921, the main equipment being No. 7½ gyratories, followed by 48-in. horizontal Symons disk crushers. Fines were removed by I-in. stationary grizzlies ahead of each unit. In 1924 a third stage of crushing was provided by the addition of 55½-in. rolls in open circuit. During 1925, the Symons disks were replaced by 73½-in. rolls. All the rolls were changed to closed circuit, with elevators and screens to give a final product of 9 pct plus 3-mesh. The flowsheet of a typical crushing section of this period is shown in Fig I. The ore was withdrawn from a common bill serving the five crushing sections in the plant by means of any number up to six 36-in. wide-apron feeders, which discharged onto a common conveyor belt. This conveyor was equipped with a magnetic head pulley and the crusher attendant was stationed near the discharge end to pick out pieces of mine timber that might be in the ore. The ore was discharged over a 13/4-in. spaced bar grizzly and the oversize of this grizzly was crushed by a No. 7½ McCully gyratory crusher set at 13/4-in. opening on the closed side. The grizzly undersize joined the gyratory discharge and was conveyed to a set of 73½ by 20-in. Garfield rolls set at I in. These rolls operated in closed circuit with a bucket elevator and two vibrating screens placed in parallel.

Jan 1, 1949