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By Thomas Fraser, H. F. Yancey

Prior to the war just ended, Germany was the greatest coal producer in continental Europe. In 1943 the production of coal, bituminous and higher in rank, together with brown coal calculated to the equivalent amount of bituminous coal, is estimated to have been about 240 million metric tons, according to the War Department. Before the war production in Germany exceeded consumption. No other country in continental Europe except Poland had a large surplus. However, England's surplus was the largest in Europe as a whole. Before World War II the Ruhr Basin normally supplied about 70 pet of the total production of high-rank coal, the Saar and Aachen about 6 pet each, Upper Silesia about 14 pet, and Silesia and Saxony the remainder. Nearly one third of the brown coal was produced in the Rhineland, a little less than one half in central Germany nest of the Elbe River, and about one fourth east of the Elbe. Bituminous coal and brown coal together constituted by far the most important energy resources of the old German Reich. Bituminous coal furnished nearly 70 pet of the total energy requirement, brown coal supplied about 25 pet, and water power, wood, petroleum and alcohol together aggregated only 7 pet. Because it was the most important single mineral resource of Germany, an excess of coal enabled the Reich to import nearly all the other mineral products that she lacked, such as iron ore from Lorraine. The area adjacent to the coal fields was supplied with gas from an interconnected grid. Similar electric grids are based on brown coal with the power stations at the mines. Coal tar provided the foundation for the development of the extensive German chemical industries as exemplified by the I. G. Farbenindustrie. For almost a century, German scientists have been adequately supported both by the State and by industry in studying improved methods of utilizing coal. In Germany the mineral rights were reserved by the State and ordered eatraction of coal was enforced. Each lease comprised a relatively extensive area. which accounts for the generally large size of the mines and mining companies In 1938 there were 160 mines in the Ruhr, with a total output of 125 million tons but 90 pet of the production was controlled by 20 companies. By far the largest single coal-mining company in Germany was the Gelsen-kirchener Bergwerke A. G., which produced nearly 27 million tons in the Ruhr in 1937. Second was Saargruben A. G. a State-owned company operating in the Saar, which produced 'lore than '3 million tons. Third was Bergwerksgesellschaft Hibernia A. G., with 10 million tons, also in the Ruhr. Briquettes are made from both bitumillous coal and brown coal, but in 1938 some 44 million toils of brown-coal bri-

Jan 1, 1947

By Thomas Fraser, H. F. Yancey

Prior to the war just ended, Germany was the greatest coal producer in continental Europe. In 1943 the production of coal, bituminous and higher in rank, together with brown coal calculated to the equivalent amount of bituminous coal, is estimated to have been about 240 million metric tons, according to the War Department. Before the war production in Germany exceeded consumption. No other country in continental Europe except Poland had a large surplus. However, England's surplus was the largest in Europe as a whole. Before World War II the Ruhr Basin normally supplied about 70 pet of the total production of high-rank coal, the Saar and Aachen about 6 pet each, Upper Silesia about 14 pet, and Silesia and Saxony the remainder. Nearly one third of the brown coal was produced in the Rhineland, a little less than one half in central Germany nest of the Elbe River, and about one fourth east of the Elbe. Bituminous coal and brown coal together constituted by far the most important energy resources of the old German Reich. Bituminous coal furnished nearly 70 pet of the total energy requirement, brown coal supplied about 25 pet, and water power, wood, petroleum and alcohol together aggregated only 7 pet. Because it was the most important single mineral resource of Germany, an excess of coal enabled the Reich to import nearly all the other mineral products that she lacked, such as iron ore from Lorraine. The area adjacent to the coal fields was supplied with gas from an interconnected grid. Similar electric grids are based on brown coal with the power stations at the mines. Coal tar provided the foundation for the development of the extensive German chemical industries as exemplified by the I. G. Farbenindustrie. For almost a century, German scientists have been adequately supported both by the State and by industry in studying improved methods of utilizing coal. In Germany the mineral rights were reserved by the State and ordered eatraction of coal was enforced. Each lease comprised a relatively extensive area. which accounts for the generally large size of the mines and mining companies In 1938 there were 160 mines in the Ruhr, with a total output of 125 million tons but 90 pet of the production was controlled by 20 companies. By far the largest single coal-mining company in Germany was the Gelsen-kirchener Bergwerke A. G., which produced nearly 27 million tons in the Ruhr in 1937. Second was Saargruben A. G. a State-owned company operating in the Saar, which produced 'lore than '3 million tons. Third was Bergwerksgesellschaft Hibernia A. G., with 10 million tons, also in the Ruhr. Briquettes are made from both bitumillous coal and brown coal, but in 1938 some 44 million toils of brown-coal bri-

Jan 1, 1947

By R. E. Miner, J. K. Jackson, T. L. Wilson

THE nature of the decomposition reactions producing hardening in aluminum-rich A1-Ag alloys is now fairly well understood. The silver is not uniformly dispersed even at temperatures above the solvus line: rather, it is clustered into zones and associated with dislocations.' During quenching, nearly all of the silver segregates into zones' and some is attracted to dislocations, producing stacking faults. Growth of these faults produces the semicoherent y precipitate.1 During aging, the larger zones and the y precipitates grow at the expense of the smaller zones. The large zones persist up to about 350°C; where they redissolve in favor of the y' and y precipitates.' The equilibrium hcp y precipitate occurs both by a cellular reaction at the grain boundaries1 and by growth of the Y' plates to a critical size where incoherency results. Damask and Nowick 4 observed an internal friction peak in an A1-6 at. pct Ag alloy at about 140°C (0.25 cps). The peak was found to exhibit a systematic instability. After aging at successively higher temperatures between 155' and 23Z°C, the peak first migrated to higher temperatures while increasing in height, then returned to lower temperatures and decreased in height. After aging at temperatures above about 200°C, the peak position remained constant while the height decreased very slowly. In quenched specimens, elastic after-effect measurements inhcated a rapid initial transient in peak position from low temperatures up to 155°C. The authors concluded that a stress induced growth and resolution of the clusters and the 7' was the most reasonable mechanism to account for the behavior. It would seem, however, very unlikely that both the clusters and the y' precipitate would provide damping contributions whose relaxation time and activation energies were identical, which is required for them to produce the same internal friction peak. The present investigation was motivated by this apparent anomaly and the development of a very low thermal inertia torsion pendulum5 capable of monitoring transient behavior. Low frequency internal friction measurements were made on four binary A1-Ag alloys nominally containing 5, 10: 15, and 20 at. pct Ag. The results of measurements made on a 10 at. pct Ag alloy in the quenched condition are shown in Fig. 1. There are two peaks apparent in the spectrum, one centered at about 160°C, the second at about 210°C. The two peak spectrum was observed in each alloy, but it was best resolved in the 10 pct Ag specimens. Because of the instability in the internal friction, 4 it seemed possible that the second peak was not a peak at all, but simply a manifestation of a single, sharply transient, peak. This effect has been observed in Au-Ni alloys.6 The existence of the higher tempera-

Jan 1, 1970

By R. W. Gurry

IN the course of a series of measurements of the rate of diffusion of carbon in austenite at about 960°C. (1760°F.) and 1110°C. (2030°F.), it became necessary to determine carbon concentration when austenite is saturated with graphite, because of indications that the hitherto accepted value of the solubility of graphite in this temperature range is too low. Specimens of carbonyl iron were saturated at temperature in a hydrogen-toluene atmosphere which was in equilibrium with graphite. The carbon content of the steel was measured either by combustion or by the loss of weight after substantially complete decarburization in an atmosphere of hydrogen saturated at room temperature with water vapor. The absence of carbide or graphite as a phase at temperature was demonstrated by microscopic examination of saturated specimens that had been drastically quenched. The mean result of concordant measurements is that gamma iron saturated with respect to graphite contains 1.39 per cent carbon at 957°C. (1755°F.) and 1.89 per cent at 1110°C. (2030°F.). These two values, plotted along with what appears from the literature to be the best value (0.69 per cent) at the iron-graphite eutectoid temperature (738°C. or 1360°F.), yield a curve that is almost linear, and which, continued upward through a short interval, leads to a limiting solubility of 1.98 per cent at the iron-graphite eutectic temperature 1135°C. (2075°F.) instead of the commonly accepted value of about 1.7 per cent. EXPERIMENTAL PROCEDURE The samples of carbonyl iron, with the exception of those for microscopic examination, were cylinders 3.8 cm. long, 0.475 cm. in diameter. They were placed in a graphite holder and suspended in a vertical furnace, the temperature of which was controlled precisely by means of the resistance of the platinum winding.1 In the empty furnace the range of temperature over a distance of 4 cm. at the heat center was 9°C. (16°F.) but at a given location the temperature remained within ±0.5°C. (0.9°F.); it was measured by inserting a platinum-rhodium couple calibrated at the melting point of palladium, 1555°C. Since the presence of the sample probably improved the uniformity of temperature, it was decided to consider the average temperature of the sample to be 3°C. lower than the maximum measured in the empty furnace with the same setting of the controller. Data were obtained in the vicinity of two temperatures, 960°C. (1760°F.) and 1110°C. (2030°F.). The samples for microscopic examination, 2.2 cm. long, 0.475 cm. in diameter, were suspended individually on pure iron wire; for these no temperature correction was applied. The carburizing gas passed into the furnace was dry, oxygen-free hydrogen saturated at room temperature with

Jan 1, 1942

By A. T. Beckwith

This paper concerns measurements of low-velocity air currents and investigations on mine ventilation by means of chemical smoke. The chemical smoke used is produced without flame and at ordinary mine temperatures, and therefore presents no ignition hazard in gaseous mines. Mining men who have worked in open-light mines are familiar with the use of tobacco smoke in determining the direction of air travel and approximately measuring low-velocity air currents. The use of chemical smoke is merely a refinement on the use of tobacco smoke and the flameless low-temperature source of chemical smoke has the additional advantage that it can be used in gaseous mines. Ventilation smoke apparatus, Fig I, is a simple device for producing dense white smoke. The apparatus as shown consists of a rubber aspirator bulb which is used to force air through a glass "smoke tube" containing a chemical substance that produces a dense white smoke on contact with air. The apparatus was developed by engineers of the U. S. Bureau of Mines.' It is inexpensive, weighs less than 3 oz, is easily carried in a pocket, and requires no skill to operate. Smoke apparatus is invaluable in practical mine ventilation work for: (I) measuring low-velocity air currents, (2) detecting air leakage, (3) observing airflow characteristics. This paper is written to show the merits of the apparatus, and is based upon findings through its extensive use, during the past six years, in mines operated by the Lehigh Navigation Coal Co. Inc. The Apparatus Certain chemicals, such as stannic chloride (SnCl4) and pyrosulphuric acid (H2S2O7) produce "white smoke" when exposed to air. Stannic chloride gives off vapors which on contact with moist air cause the formation of solid hydrates, of which SnCl4.5H20 is the most important. Pyrosulphuric acid decomposes when exposed to air with the evolution of sulphur trioxide (SO3) fumes, which become white upon contacting moisture in the air. A chemical, such as one of the above, is the active substance in a smoke tube. The glass tubes are hermetically sealed during manufacture, and are made ready for use by merely breaking off the tips at the conveniently located file cuts. When the smoke tube is connected to the aspirator, depressing the aspirator bulb forces air into contact with the chemical, and smoke discharges from the open end of the tube. A white solid (a hydrate if stannic chloride is used, or sulphur trioxide if pyrosulphuric acid is used) may occasionally form a cake in the outlet opening of a smoke tube. A small wire such as an ordinary paper clip can be conveniently employed to dislodge the cake should it seal the tube. When the assembled apparatus is not

Jan 1, 1949

By Henry S. Drinker

THE Butler Mine property is situated in the vicinity of Pittston, in the Wyoming coal-field of Pennsylvania. The coal has been worked out from the fourteen-foot or Baltimore vein for a number of years on part of the tract, the old chambers remaining open as when originally abandoned. This vein outcrops on the Butler property, and is nowhere more than sixty feet below the surface of the ground on the line of the cut-off to be described. Above the fourteen-foot vein and about, say midway, between it and the surface of the ground, there is another small vein or rider of coal, about twelve inches in thickness. The rock between the two veins is a carbonaceous slate, about twenty-five feet in thickness. Above the smaller vein is a bed of sandstone, varying in thickness, but generally not less than from eight to ten feet. This sandstone is also carbonaceous, but the proportion of carbon is much less than in the slate below. Both rocks have been analyzed, and before the next Institute meeting, the writer hopes to be able to furnish the exact proportions. The sandstone is overlaid with debris up to the surface of the ground. Some time during the spring of 1876, it was found that the coal left in the worked-out mine had been set on fire, and that the fire was gradually spreading. Inquiry developed the fact that the fire had originally been started by a woman, who had taken up her abode in an old tunnel leading into the workings. Owing to some delay in the early attempts made to extinguish the fire, it finally gained such headway that no ordinary means were found to be effectual. Finally, Mr. C. F. Conrad, civil and mining engineer, of Pittston, Penna., was called in as consulting engineer. He advised the making of a thorough cut through such portion of the old workings as had not been reached by the fire. It was found by survey that this cut could be limited to a length of about one thousand two hundred feet, as the vein faulted on both sides, and that the greatest depth of the cut would be about sixty feet. Subsequently it was decided to tunnel part of the way, thus saving the deepest portion of the cut. In 1856-7, there occurred a fire in the same mine. Two stone-walls in V-shape running from the apex of the V towards the outcrop, were built to cut the fire off, and these walls were carried as far as the vein had at that time been worked. After the walls were finished

Jan 1, 1879

By Henry S. Drinker

The Butler Mine property is situated in the vicinity of Pittston, in the Wyoming coal-field of Pennsylvania. The coal has been ' worked out from the fourteen-foot or Baltimore vein for a number of years on part of the tract, the old chambers remaining open as when originally abandoned. This vein outcrops on the Butler prop erty, and is nowhere more than sixty feet below the surface of the ground on the line of the cut-off to be described. Above the four teen-foot vein and about, say midway, between it and the surface of the ground, there is another small vein or rider of coal, about twelve inches in thickness. The rock between the two veins is a carbon aceous slate, about twenty-five feet in thickness. Above the smaller vein is a bed of sandstone, varying in thickness, but generally not less than from eight to ten feet. This sandstone is also carbonaceous, but the proportion of carbon is much less than in the slate below. Both rocks have been analyzed, and before the next Institute meet ing, the writer hopes to be able to furnish the exact proportions. The sandstone is overlaid with debris up to the surface of the ground. Some time during the spring of 1876, it was found that the coal left in the worked-out mine had been set on fire, and that the fire was gradually spreading. Inquiry developed the fact that the fire had originally been started by a woman, who had taken up her abode in an old tunnel leading into the workings. Owing to some delay in the early attempts made to extinguish the fire, it finally gained such headway that no ordinary means were found to be effectual. Finally, Mr. C. F. Conrad, civil and mining engineer, of Pittston, Penna., was called in as consulting engineer. He advised the making of a thorough cut through such portion of the old workings as had not been reached by the fire. It was found by survey that this cut could be limited to a length of about one thousand two hundred feet, as the vein faulted on both sides, and that the greatest depth of the cut would be about sixty feet. Subsequently it was decided to tunnel part of the way, thus saving the deepest portion of the cut. In 1856-7, there occurred a fire in the same mine. Two stone walls in V-shape running from the apex of the V towards the out crop, were built to cot the fire 06 and these walls were carried as far as the vein had at that time been worked. After the walls were finished

Jan 1, 1879

By Thomas F. Donnelly

Discussion of the paper of THOMAS F. DONNELLY, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 104, August, 1915, pp. 1759 to 1768. F. LYNWOOD GARRISON, Philadelphia, Pa. (communication to the Secretary *).-The paper is of value in giving the public some first-hand information regarding a copper property that from various causes has attracted considerable attention, especially in Philadelphia, and which has also been the source of several fantastic flotations that had little to recommend them. Since 1906, no work has been done on this property under my direction, with the exception of a topographic survey made last winter by Mr. Donnelly, which was intended as the first step in a sound and logical system of development. From 1906 to 1915, desultory attempts were made to develop these orebodies, but as the work was not in competent hands much money was wasted and even more disappeared in ways regarding which it is perhaps not best to inquire. The result is, this interesting property has now practically no more ore developed upon it than in 1906, despite the fact that a new 100-ton mill is being erected, although, as I have intimated, it is today not much more than a prospect. Turning now to the scientific and more agreeable side of the subject, it is interesting to note that these Santo Domingo copper ores are invariably found in volcanic tuffs and that, in a geologic sense, they are all very old, for there are no evidences of recent volcanic activity in the greater Antilles, of which this island is the second largest. The existence of copper in tuffs in different parts of the world is a notable fact. Often, if not usually, it is present in mere traces, but not infrequently we find it in sufficient amount to be of economic importance. Thus, for example, chalcopyrite and bornite occur in the Tertiary tuffs of Japan, and a number of producing copper mines of that country are either actually in, or closely associated with, rocks of this character. Whether the copper minerals are original primary ores in the tuffs, or were derived from accompanying eruptive dikes, appears not yet to have been determined in any instance. In most of the Japanese cases of which I can find any record the tuffs appear to be associated with shales and limestones. These Japanese copper ores also carry some gold, especially in the upper parts of the veins; in this they resemble the Santo Domingo deposits which not infrequently carry as much as $6 or $7 per ton of gold on the outcrop. As a rule the content of gold in both the Japanese and

Jan 12, 1915

JOHN W. BELL (communication to the Secretary*).-I have to thank Mr. Merrill not only for his very kind remarks, but for the weight of his implied opinion that the work accomplished probably has an important practical aspect. I think there is a distinct although unuttered impression among engineers, that it will be time enough for them to begin to take an interest in the law of crushing when it emerges from the cloud of violent and even vituperative controversy that has hovered over it during the last 6 or 7 years. I have a very confident feeling that as the result of the work done by Mr. Gates, by me and by Messrs. Mitchell, Baily and Cockfield, who so loyally helped me to obtain the necessary data, at least the practical law of crushing has emerged; that it is ready and available for use by our fellow engineers, and that it will help them to unravel a number of the many practical problems they are con¬stantly confronted with. In regard to Mr. Gates' discussion, I would indeed like to believe that McGill could claim, as a minor achievement, the credit which Mr. Gates so generously bestows of having established the "law" of crushing. I think that more work would be required to establish this claim in the, strict sense of the word, as there is a fairly large gap between the coarse and fine crushing results. I am not worried about that gap, because the reason for it is quite clear to me; at the same time I-think it will be well to bridge it by a series of tests, and I am bound to admit that the testing machine is the most promising machine right at hand to work with, although whether the minute amounts of power required to crush small quantities of rock can be so accurately integrated as to insure high accuracy in the final result is still a question in my mind. If we had an Amsler-Laffon machine I would certainly try it. Mr. Gates' remarks about choke crushing are extremely interesting. He compels me to wonder, too. It is a new and probably a very important factor to be considered in this connection. I am in perfect accord with Mr. Gates about the desirability of the determination of crushing constants, because when this is accomplished the efficiencies (not relative mechanical, but real efficiencies) of different types of rock-crushing plants in any part of the world will then be comparable. For this reason it would, seem desirable that the American Institute Committee should not only consider ways and means for accom-

Jan 6, 1917

By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

By R. L. Featherly, J. R. James

The replacement or reconditioning of oil storage tank bottoms due to external corrosion is an expensive maintenance item to the oil industry. Cathodic protection as a means of mitigating this problem has proved very successful in the past few years. One pipe-line company recently used magnesium anodes to cathodically protect two 55,000 bbl and two 20,000 bbl tanks. The tank bottoms were protected as individual units with considerable factual data being taken on one of the 20,000 bbl tanks. Twelve 60-lb anodes were installed about 2½ ft from the tank on approximately 17-½ft centers around the tank. Previous installations and tests have shown that sufficient protection is provided when the tank-to-soil potential using a copper-copper sulphate half-cell as a reference electrode is — 0.90 volts at the edge of tank. Potential measurements taken a few months after completing the installation indicated that the potentials were higher than were needed for protection. Resistors were then inserted in the individual anode circuits which increased the life of the anode installation and still provided adequate protection for the structure. All connecting pipe lines were then insulated from the tank. This limited the anode current strictly to the tank resulting in a greater degree of protection with a still longer anticipated anode installation life. The total installation cost for this tank was less than $300 which amounts to $20 per year The corrosion of metallic structures in contact with soil or water has been a very serious and costly problem for centuries. The annual loss to our nation's pipelines alone is estimated at $200,000,000. Mitigation of this problem by various methods has been carried on quite successfully for a number of years. An electrical method called cathodic protection has been used with considerable success to control corrosive influences on pipelines, tank bottoms, cable sheathing and other buried metal structures. Corrosion of oil storage tank bottoms has been a serious problem among the oil industries for years. In addition to damage costs and loss of oil, the replacement of one tank alone represents several thousand dollars. A reconditioning program at periodic intervals was instituted by several organizations and found to be a very expensive process. A few years ago two of the pipeline companies installed cathodic protection units at their tank farms.3'4 The average current required for protection of the tank bottoms varied from 0.7 to 1.0 milliampere per square foot. The results indicated that adequate protection was obtained at a cost of approximately 10 pct of the previous reconditioning program. The Republic Pipe Line Company experienced severe corrosion on their oil storage tanks. One of their 55,000-bbl tanks

Jan 1, 1948

By R. L. Featherly, J. R. James

The replacement or reconditioning of oil storage tank bottoms due to external corrosion is an expensive maintenance item to the oil industry. Cathodic protection as a means of mitigating this problem has proved very successful in the past few years. One pipe-line company recently used magnesium anodes to cathodically protect two 55,000 bbl and two 20,000 bbl tanks. The tank bottoms were protected as individual units with considerable factual data being taken on one of the 20,000 bbl tanks. Twelve 60-lb anodes were installed about 2½ ft from the tank on approximately 17-½ft centers around the tank. Previous installations and tests have shown that sufficient protection is provided when the tank-to-soil potential using a copper-copper sulphate half-cell as a reference electrode is — 0.90 volts at the edge of tank. Potential measurements taken a few months after completing the installation indicated that the potentials were higher than were needed for protection. Resistors were then inserted in the individual anode circuits which increased the life of the anode installation and still provided adequate protection for the structure. All connecting pipe lines were then insulated from the tank. This limited the anode current strictly to the tank resulting in a greater degree of protection with a still longer anticipated anode installation life. The total installation cost for this tank was less than $300 which amounts to $20 per year The corrosion of metallic structures in contact with soil or water has been a very serious and costly problem for centuries. The annual loss to our nation's pipelines alone is estimated at $200,000,000. Mitigation of this problem by various methods has been carried on quite successfully for a number of years. An electrical method called cathodic protection has been used with considerable success to control corrosive influences on pipelines, tank bottoms, cable sheathing and other buried metal structures. Corrosion of oil storage tank bottoms has been a serious problem among the oil industries for years. In addition to damage costs and loss of oil, the replacement of one tank alone represents several thousand dollars. A reconditioning program at periodic intervals was instituted by several organizations and found to be a very expensive process. A few years ago two of the pipeline companies installed cathodic protection units at their tank farms.3'4 The average current required for protection of the tank bottoms varied from 0.7 to 1.0 milliampere per square foot. The results indicated that adequate protection was obtained at a cost of approximately 10 pct of the previous reconditioning program. The Republic Pipe Line Company experienced severe corrosion on their oil storage tanks. One of their 55,000-bbl tanks

Jan 1, 1948

The Gooseberry Mine is located 38 krn (24 miles) by road east of Reno, 13 km (8 miles) south of Interstate 80. The mine was originally discovered in 1906 and remained in the control of the Martin family of Fresno, who worked the-property with Mexican ranch hands for 68 years. Westcoast took over the property in 1976 from Apco Oil Corporation. During the time the Martins operated the mine, a 305 m (1,000 ft) inclined shaft was sunk with drifts developed on eight levels exceeding a total of 3,000 m (9,000 ft) in length. It is interesting to note that for over one -half century all development ore was stockpiled. West- coast, the United States mining subsidiary of Home Oil Company Ltd., Calgary, Alberta, completed sinking another vertical shaft to the 396 m (1,300 ft) level. Development work between 122 m and 305 m (400 ft and 1,000 ft) levels has proved the existence of a fine ore body over 300 m (1,000 ft) long. The silver-gold ore grade is fairly consistent in the vertical section, but copper, lead and zinc may show slight increases with depth. Average width of the ore bearing portion of the vein is 2.5 m (8.3 ft). Stoping is by cut-and-fill method. Average silver values within the ore zone have increased from the 152 m (500 ft) level to the 305 m (1,000 ft) level, although the gold values do show some decrease. It is not possible for the geological staff to predict the depth of the ore. However, several years of ore reserves have been proved in the existing mining areas. Important ore minerals are argentite, native gold, and electrum. Trace amounts of galena, chalcopyrite, and sphalerite are associated with the precious metal mineralization. The vein is mainly composed of calcite and quartz. Mine-run underground ore is dumped from the hoisting skip into a coarse ore bin followed by a jaw crusher, an impact crusher, and a 600 mt (650 st) fine ore bin. The mill has a capacity of 317 mt per day (350 st per day). Two-stage grinding is done by a rod mill followed by cyclones with the underflow sent to a ball mill in closed circuit with the cyclones. Cyclone overflow after conditioning receives a primary float in four 1.7 cu m (30 cu ft) cells and a second stage of conditioning ahead of 24 0.8 cu m (30 cu ft) secondary flotation cells. No cleaning stages are used. Flotation tailings are cycloned for a sand-slime separation with the sands used as underground stope fill and the slime portion to tailing dam storage.

Jan 1, 1981

By Bunting S. Crocker

DISCUSSION Fred C. Bond (Allis-Chalmers Manufacturing Co., Milwaukee)—This is a very complete description of the screened ore fine grinding installation at Lake Shore. The only important data which might be desired are the operating screen analyses and power consumption in each stage. The statement that the energy consumption, or kw-hr per ton, remains constant for all mill diameters is of great interest, but the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases. More definite test work is needed to establish this point. Screened ore fine minding has been successful on hard Lake Shore ore &d on about equally hard South African Rand ore. However, it is still not determined whether it would be successful on ores of average or less than average hardness. It is possible that with softer ores the consumption of ore pebbles might be exceedingly high, or might even exceed the available supply. Also, the softer feed might result in a rapid buildup of harmful intermediate sizes in the mill circuit. These factors can be determined only by a full scale or long pilot mill test. In any case the Lake Shore staff is to be congratulated for this pioneering work. Bunting. S. Crocker (author's reply)—As requested, the operating screen analyses and Bower consumption for each of the four stages of pebble milling are given, see Table XII. The fist three stages are in series circuit; the last stage is in closed circuit. The primary circuit had rock feed averaging 124 g per piece and the final stages rock feed weighing 63 g. The intermediate stages had mixtures of the two. Regarding Mr. Bond's statement that "the opinion is widely held that the kw-hr per ton decreases as the mill diameter increases," I would say that I have frequently been told this but have never yet been shown any evidence or seen any evidence published that would prove this statement. It may be true that some of the big mills recently installed have mechanical power savers such as micarta bearings and better lubrication, but as far as the power consumed inside the mill is concerned, we at Lake Shore think that it varies as D2 .", and that is all. We published extensive information on this point in our 1939 paper. Recently Mr. Banks at Consolidated published the figures on his 10 ft 8 in. x 12 ft 0 in. rod mill. As reported in this discussion by the writer: we calculated the power necessary to run this mill based on our 7x6-ft rod mill to within 1 pct of the actual power reported by Mr.

Jan 1, 1953

By C. M. Romanowitz, H. A. Sawin

PLACER dredging as we know it today, especially gold dredging, is an industry about 40 years old, dating from the beginning of this century, when a few mining men in California saw the possibilities in adapting earlier dredging methods to California placer fields in and immediately adjacent to streams. At first it was thought that dredges must be operated in existing stream beds, but experience soon pointed the way to dredging alluvial deposits of old stream beds, operating on landlocked dredging ponds. Within the first five or six years of modern dredging, experience was gained so rapidly that dredges to dig 60 ft. below water level were designed and built. This depth was believed by many operators to be the maximum at which successful dredging could be carried on. Alluvial deposits of gold, suitable for dredging, were readily handled by equipment available and developed early in the industry's history. Much of the gravel was easily dug and if deposits were cemented, contained large boulders, or for some other reason could not be handled by a dredge, the ground was abandoned in favor of other deposits that could be handled easily. Many failures, adversely affecting the dredging industry, resulted from attempts at that time to extend the scope of placer dredging with equipment not suited to the formation being dredged. Dredging ground suited to early equipment was exhausted eventually, but many supposedly undredgeable alluvial fields containing gold in commercial quantities were found and became known to dredgemen. Improved equipment was needed to work deposits known to exist at great depths or associated with unusual obstructions. Operating companies looked to dredge manufacturers and metallurgists, particularly manganese-steel foundries, for improved designs and materials. Experience developed in California has spread throughout the world, and dredges of improved types have been adapted to the mining of platinum and tin as well as gold. Space will not permit a description of all the improvements in dredging since the early days, but a few of the recent developments are worthy of note, because they make possible the dredging of placer deposits of great depth, those having hard, irregular bedrock or containing large boulders in great quantities, and compacted or cemented gravel deposits. DEEP DREDGING At Hammonton, Calif., Yuba Consolidated Gold Fields has, within the past two years, built its dredge No. 20. The company has operated continuously in that field since 1904, digging successively at maximum depths of 60 ft., 80 ft., 112 ft., and now 124 ft. below water level. The first dredges were equipped with buckets of 6 and 7 1/2-cu. ft. capacity; six dredges operating in the field today have 18-cu. ft. buckets, digging now 65 to 124 ft. below water level. Some ground near Hammonton has been dug for the third time, owing to the devel-

Jan 1, 1941

By Richard M. Foose

FIVE miles southwest of Mt. Holly Springs, Cumberland County, Pennsylvania, the Philadelphia Clay Co. has been mining and milling a white clay since 1896; for use in white cement, as a filler in rubber, pigment and paper, in the manufacture of insecticides, and for other minor purposes. Until late in 1942, when the mine was abandoned because of labor shortage, combined underground mining and surface stripping produced approximately 30,000 tons of clay a year, of which 10,000 tons were refined in the mill. GEOLOGY Lower Cambrian Antietam quartzite and the Montalto quartzite, stratigraphically older, trend N.55°E. and make a sharp 1600-ft. anticlinal ridge of the South Mountains. The Tomstown dolomite overlies the Antietam quartzite on the southeast flank of the anticline, which is displaced near its crest by a steeply dipping oblique fault. Erosion of the less resistant Tomstown formation has resulted in a narrow valley at an elevation of 700 ft. Directly on top of the Antietam quartzite and at the base of the Tomstown formation is a zone of sandy white clay, lenticular along the strike of the formations, ranging in thickness from 10 to 300 ft. within a mile. All of the formations dip steeply to the southeast, from 65° to vertical. The clay crops out at an elevation of about 1000 feet. ORIGIN AND CHARACTER The uppermost part of the Antietam formation is sericitic and schistose at many places, and at all these places weathering is in an advanced stage. The clay is the residual product of weathering of this sericitic quartzite or sandy schist. The crude clay has the physical aspect of a sandy kaolin, but there are large areas in the clay formation that are low in free silica, and commonly the 30 ft. adjacent to the Antietam quartzite is notably free of silica. The overlying Tomstown dolomite has been weathered almost completely to a mixed yellow and brown clay, which carries a zone of nodular and chunky limonite and manganese ore, 4 to 12 ft. thick, about l00 ft. from the contact with the white clay. In mining, this zone is an important "marker," indicating where the white clay may be expected in an advancing tunnel. The contact of the white clay formation with the yellow and brown clay of the Tomstown formation is relatively sharp, although the line of contact is irregular and therefore responsible for the lenticular character of the white clay formation. Following is an analysis of the refined clay containing less than one per cent of moisture when dried at 212°F.: Si02, 67.05 per cent; A1203, 22.27; Fe203, 0.40; CaO, 0.00; Ti02, 1.05; MgO, 0.52; Na20, 0.04; K20, 0.88; ignition loss, 7.65; total, 99.86.

Jan 1, 1944

By A. Kenneth Schellinger, O. Cutler Shepard

INTRODUCTION OVERGRINDING, or the breaking of ore particles into sizes smaller than required for liberation, is a first-magnitude problem in grinding for concentration processes. The conventional ball mill-classifier circuit used in commercial grinding produces mineral particles that are all ground to pass an upper size limit, while the proportion of very fine sizes, or slimes, is uncontrolled. Calculations based on measurements made in commercial plants which were grinding ore for flotation concentration indicate that about 90 pct of the energy used in grinding is consumed in useless overgrinding.1 In addition to accounting for the consumption of a large amount of energy, overgrinding is detrimental to the flotation process. Overground slime coats larger particles and inhibits selective action by the flotation reagents on particles which by themselves could be separated very well. A further difficulty is that extremely finely ground minerals do not respond satisfactorily to flotation-collecting reagents, and both selection and recovery decrease markedly with particles finer than 5 to 10 microns.2 Conventional classifier-ball mill circuits inherently tend to overgrind the valuable minerals more than the gangue minerals because most valuable minerals are of higher density and are softer and more friable than the gangue minerals. The remedy to this situation would seem to be a separating mechanism which would separate free valuable mineral particles regardless of size and density considerations and a mechanism which would operate in the grinding mill so that valuable mineral grains would be removed from the grinding machine as soon as they are freed. A consideration of these ideas led the authors to the conception of a simultaneous grinding and flotation mechanism. So far as the authors are aware, such a machine has never been tried, even on a laboratory scale. DEVELOPMENT OF APPARATUS AND EXPERIMENTAL METHOD Apparatus development went through several stages before a really workable machine was devised. Work was started with a large cast-iron mortar 7 ½ in. in diameter and 8 in. deep. A cast-iron disk was mounted on the end of a vertical shaft so that it fitted loosely into the bottom of the mortar. On top of the disk four radial cleats ¼ in. high were bolted to agitate the ball charge as the disk, or impeller, rotated. The vertical shaft was driven at 100 rpm by a mechanism improvised from an old coffeemill type of grinder. A hole was drilled into the bottom of the mortar through which compressed air was introduced for frothing. In operation, weighed charges of ore which had been crushed to pass an 8-mesh screen were placed in the bowl along with water and suitable reagents. The impeller was then started and the compressed air turned on so that the ore was ground while being exposed to air bubbles. Mineralized froth was

Jan 1, 1947

By J. G. Giddings, R. Kobayashi

Residual viscosity, the ViSCOSitY at a given pressure and temperature minus the dilute gas viscosity at the same temperature, has been found to be independent of temperature for pure components and mixtures for simple molecules and light hydrocarbons. The residual viscosities for methane, ethane, propane and four mixtures were correlated as a function of reduced density and molecular weight. INTRODUCTION Earlier efforts in correlating the viscosity of light hydrocarbons and associated gases under pressure, primarily by some form of the corresponding states principle, have been reported elsewhere.7 The formal theory of corresponding states principle for viscosity from non-equilibrium statistical mechanical arguments for simple molecules has been developed by Helfand and Rice.13 The most serious obstacle to the development of a generalized corresponding states representation of viscosity data, even for simple molecules, has been the limited quantity of precise viscosity data of pure components and mixtures over a wide range of pressure and temperature conditions. Viscosity data on light hydrocarbons have recently been contributed by several laboratories.2,6,8,23,25,26 In the study of the viscosity of nitrogen, Michels and Gibson" observed that the viscosities of nitrogen as a function of density gave parallel isotherms. The same behavior has since been reported for other gases.".= The observed behavior indicates that the residual viscosity or the viscosity at a given pressure and temperature minus the dilute gas viscosity at the same temperature is a unique function of the density. Abas-Zade,1 Michels and Botzen,18 and Michels, Bot-Zen, Friedman and Sengers,19 observed similar behavior for thermal conductivity, the residual thermal conductivity being defined in a manner parallel to that of residual viscosity. Thodos and co-workers 4,15,24 applied the residual viscosity concept to the viscosity of pure monatomic and diatomic gases. Ellington and co-workers10,25 made the residual viscosity vs density plots for ethane and propane, but omitted the region below a residual viscosity of 15 micropoise and showed a systematic separation of the ethane isotherms in the low density region. Eakin correlated the residual viscosity of methane, ethane, propane and n-butane empirically as a function of density and molecular weight: Carmichael and Sage6 recently presented ethane residual viscosity vs density values in the gaseous liquid, and near critical regions as a single univariant function of density. Flynn, Hanks, Lemaire and ROSS" have contributed extremely accurate viscosity data to confirm that the residual viscosity concept is essentially correct, but may not be accurate within 0.1 per cent of the viscosity value for the gaseous state- The recent studies by Giddings12 indicate that the residual viscosity of methane, propane and each of four methane-propane mixtures are unique functions of density Over the temperature and the density ranges studied.

Jan 1, 1965

By Tell Ertl, Ernest E. Burgh

The Bureau of Mines is operating an oil-shale mine 10 miles west of Rifle, Colo., as part of its Synthetic Liquid Fuels program. The purpose of operating this mine is twofold: First, to supply oil shale to an experimental Processing plant; and second, to develop practices and select methods for mining oil shale on a commercial scale at the lowest practicable cost. The second purpose—that of finding out how to mine oil shale most cheaply—entails considerable mining research. Since the drilling of oil shale probably will be one of the major items of the mining cost, fattors that influence drilling speeds were investigated. One phase of the drilling program was the development of a successful hard-surfaced rock-drill bit1 that can be used for drilling long holes in oil shale without changing bits. During that phase, drilling speeds were measured and drill cuttings were collected for assay and for screen sizing, when the cuttings from the individual tests were compared, a definite relationship was shown to exist between the drilling speed and the coarseness of the drill cuttings. Procedure Measurement of Drilling Speeds All drilling in the tests was done with a Model UMB-D-99W Gardner-Denver wagon drill. Round-lugged drill steel, 141 in. in diameter was used in 6, 12, and 18 ft lengths with 1/8-in. Timken H-type bits hard-faced with acetylene tube borium. The tests were made in the lower adit of the Bureau of Mines oil-shale mine; the grade of the oil shale drilled ranged from 18.6 to 42.7 gal per ton, During the experiments air pressure varied between 75 and go psi while water pressure remained fairly constant at 75 psi. All holes were drilled almost horizontally. Each drill hole was thoroughly cleaned of cuttings after being collared. The advance of the drill along the carriage was measured and recorded as the drilling speed in inches per minute. Method of Collecting Cuttings TO facilitate the collection of cuttings, each drill hole was collared in a smooth, vertical portion of the oil-shale breast (Fig 1). A galvanized-stee1 sheet was fastened to the breast below the area in which the hole was to be drilled. The cuttings washed out of the drill hole by the drilling water were guided by the steel sheet into a wooden box. The cuttings that remained in the drill hole after the drill was withdrawn were scraped out and placed in the same box as the cuttings washed out of the hole. After the cuttings from a I-min. drilling interval were collected, they were placed in flat metal pans. The fines were allowed to settle and the clear water was decanted. The cuttings then were dried in the sun.

Jan 1, 1949

By John E. Conley

THE slate industry of the United States has shown a marked decline in value of products made annually since the peak year 1925, although there has been moderate improvement over the lean years 1932 to 1934. In 1940, the value of all slate products sold was approximately 5 3/4 million dollars for a total output of 473,450 short tons. The quantity sold was a little less and the value a great deal lower than the annual averages for the past two decades. Enormous quantities of waste materials result from the quarrying, splitting and milling of slate. Obviously, utilization of such waste material is desirable, not only from the viewpoint of economical operation but also for the purpose of conserving natural resources. One possible use is to produce expanded or puffed lightweight aggregate for use in concretes. Such concretes are strong enough for specific uses, and the advantage of lightness could be obtained in such structures as the walls and roofs of buildings, suspension-bridge roadways and various other applications where weight is a factor. Conceivably, these uses could consume an appreciable annual tonnage. USE OF WASTE SLATES FOR MAKING AGGREGATES The lightweight-aggregate industry has been discussed elsewhere4.5 and need not be covered in detail in this paper. The materials commonly utilized are naturally occurring aggregates such as volcanic cinder, tuff and pumice; by-product materials, such as ordinary cinders and air-cooled slags; and the specially manufactured products, including Haydite, Pottsco, Gravelite, Cel-seal, Superock and Waylite. Slates, however, have not been given the consideration they merit, especially in the United States. In the British Isles, some work has been done' on certain Welsh slates by the Building Research Station to stimulate the use of these wastes. REQUIREMENTS FOR RAW MATERIALS As pointed out by Hughes,4 the lightweight-aggregate industry is segregated in regions adjacent to large cities in order to localize the market and avoid the necessity of long freight haulage. Waste slate from the deposits of eastern Pennsylvania, northern Maryland, and Virginia would meet the basic requirement of any low-priced structural raw material. One apparent advantage is attained when slate waste, normally discarded, is utilized in the manufacture of a by-product. Mining or quarrying costs usually have been assessed against the main product. Some additional grinding might be desirable, but this operation would be relatively cheap, as most slates grind easily. It is believed that virtually all slates, when properly heated, will expand to some degree. However, types that will expand from 3 to 7 times their original volume are preferred. Fig. I illustrates the expansion of a Pennsylvania slate. The cube originally

Jan 1, 1942