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By W. J. Sharwood

(San Francisco Meeting, September, 1915) CUPELLATION is well known to be one of the most effective methods of separating silver and gold from base metals and other impurities, as well as one of the most accurate means for their estimation. In the latter application it consists in absorbing in some porous medium the fused lead. oxide formed by the oxidation of the lead "bath," which carries with it the oxides of other base metals present. It has long been recognized that small amounts of both silver and gold were removed and absorbed together with the base metals, as well as being to some slight extent volatilized, and in work of the highest accuracy a correction is made for the "cupellation loss." Thus. in the assay of bullion this is done by means of "proofs" of like composition, which are manipulated throughout in precisely the same manner as the assay pieces, and are assumed to lose or gain like amounts. In the case of certain very rich ores and commercial products the loss is adjusted by assaying the cupel itself and recovering the precious metal absorbed, but this doubles the labor and does not remedy the loss by volatilization. Arbitrary tables have been drawn up by various authorities with the object of enabling a correction to be applied without the use of a check. These have been of very limited practical use owing to the fact that a number of factors affect the losses sustained. For a fixed amount of precious metal the loss varies with the amount of lead used, with the nature and amount of the impurities, with the porosity of the cupel, with the air supply, and above all with the temperature at which cupellation is carried on. Pure silver loses relatively much more than gold, but the loss in either is diminished by au addition of the other metal. The concentration of precious metal in the litharge increases as the concentration in the lead increases, but little work has been published connecting the two. If all other conditions remain the same the actual total loss increases, but the percentage loss diminishes, with an increase in the weight of precious metal treated.

Jan 8, 1915

By Vincent Gianella

Barite deposits are of widespread occurrence in Nevada but there are few pro-ducing properties; most of the latter are in northern Nevada. The production of the state is small at present-in the neighborhood of 15,000 tons annually-but is steadily increasing. The barite occurs in veins and also as replacement deposits in lime-stone. The bulk of the production has come from the replacement deposits. It is thought that the barite was deposited from rising hot waters similar to-and probably contemporaneous with-those responsible for some of the metallic ores of the state. This paper discusses only the deposits in northern Nevada. THE commercial mining of barite did not begin in Nevada until 1916 but although the production has been small there has been a rapid increase in tonnage during the past several years. Barite is widely distributed in the state and production comes from at least nine of the seventeen counties. Apparently this is not generally known, as only two counties are mentioned in a recent publications dealing with barite. The increase in output is due largely to recently discovered barite deposits in northern Nevada. The principal production has come from the northern portions of Lander and Eureka counties; smaller amounts from Elko County. The discussion in this paper is confined largely to the deposits in these counties. Most of the producing area is tributary to the Western Pacific and Southern Pacific Railroads, whose tracks lie along the Humboldt River. Barite is known from other northern counties but so far the production has been small. The output has come entirely from veins and replacement deposits, the latter largely in sedimentary rocks. In general, the barite from veins is white; that from the replacement deposits usually is dark gray to black, although some is light gray or white. Most of the veins contain sulphides and many barite-bearing veins have been worked for their content of base or precious metals.

Jan 1, 1940

By Pual M. Tyler

BLIZKRIEG tactics in the present war have consumed metals on such a profligate scale that some of the best-laid procurement plans for civilian and military needs of even a year ago seem in retrospect incredibly meager. Conservation of raw materials has reached an all-time climax in Germany, where metals are worth more than human life and throwing away an empty toothpaste tube may be legally punished by death. Even the United States and the British Empire, proud masters of nine tenths of the world?s known mineral reserves and processing plants, are being driven to metals and feverishly expanding production facilities that not long ago were deemed overbuilt. After the war is over and inevitable postwar readjustments are completed, it is reasonable to expect that purchasing power ? or at least the ability to procure the comforts and luxuries of life ? will be distributed among vastly larger population than ever before, accelerating the rising trend in demand for industrial raw materials in ever greater quantity and variety.

Jan 1, 1942

By G. A. Vissac

O. R. LYONS *—I wish to thank Mr. Vissac for his compliment. I hope that his paper is not only well received, but that it will serve to bring forth more papers on the subject of thermal drying. One of the primary purposes of the work performed by Battelle for Bituminous Coal Research in investigating the thermal drying of coal was to stimulate other investigators and to get them to contribute their knowledge in the form of papers such as this one. We at Battelle and the personnel of Bituminous Coal Research are very gratified that Mr. Vissac and other persons have responded in this matter of the thermal drying of coal. I wish to state that I think that Mr. Vissac's paper is a very clear and easily understood description of a method of calculating the design requirements for a screen type drier, and I think that it would be exceedingly valuable to operators and to those who intend to purchase any type of thermal drier and use it in the future, if the manufacturers or operators who have such information for other types of driers would provide the same type of information for the other makes of driers now on the market. 1 also wish to point out—an idea that is new to me, and I know is new to most of the operators of driers in the United States-—the idea of recovering the heat that is normally lost in the coal and in the exhaust gases. This heat is not being recovered at most (of the thermal drying operations in the United States, and the possibility of recovering it should be called to the attention of every single one of those operators. I know many of them have never given any thought to the matter, but they will be interested once they realize the ease with which it could be done and the savings that could be realized. I also wish to compliment Mr. Vissac for presenting the method of analysis that he uses to determine the difficulty of drying any particular coal. It is a very simple method, and yet it seems to me that it should be a very effective, very efficient method for determining the difficulty of drying for his particular problems. C. Y. HEINER*—I do not know that I can add anything very illuminating to what Mr. Vissac has said. I think anything that Mr. Vissac said in regard to coal drying is a contribution because, to my personal knowledge, he has studied the matter carefully for many years and made many valuable contributions. I am not too familiar with coal drying problems in the east, but I know in the west we have not made enough coal drying studies. I think coal operators too often just take the coal as it is and make more or less the best of it. There are relatively few washing plants in the west now, and so the problem has not come to the front as much as it probably will in the future. In this connection, it seems to me that this matter of drying the raw coal, as Mr. Vissac brings up, is an extremely important one. We have not a continuous miner ourselves, yet, but we expect to get some this year, and we think the percentage of fine coal-—that is, minus 3/16 in.—will double. We have about 20 pct minus 3/16 in. in the 8 in. by 0 size now, and we think we will likely have 40 pct, which will have a surface moisture of the order of 8 pct. To wash it satisfactorily, we will have to dry the raw coal first in order to screen it, and after that, I suppose, there will have to be dry cleaning of some sort. We have not really used dry cleaning on fines in the west yet to my knowledge, but it is a matter that has to be faced by the industry, and I am very hopeful that Mr. Vissac's study will assist us in that connection. W. L. McMORRIS*-In my company we are preparing largely metallurgical coal for a great number of byproduct coke plants. The most outstanding thing to me about the requirements of moisture in the finished product is that there is a different requirement for almost every coke plant. Each operator has a different set of factors on which he establishes his coking costs where they involve moisture. For our corporation operations in Birmingham, my company does not produce the coal, but in Birmingham they are getting away with moistures very much higher than our plant at Clairton, Pa., would tolerate. The moisture that we have to produce for the plants along the lakefront where they are subject to much more severe weather is something else again. We have not tackled heat drying, primarily because our customers do not know what heat drying will do to the coking characteristics of the coal. If the temperature of drying can be held down

Jan 1, 1950

By AIME AIME

NEARLY two hundred 'engineers attended the dinner given at the Hotel Pennsylvania on Monday-evening, Oct. 10, to the delegation from the American l3nginiering Societies to Great Britain and France when the John Fritz Medal was presented to Sir Robert Hadfield and to Eugene Schneider. Representatives of the British and French Governments ' and of foreign engineering societies gave to the event an international aspect unique in gatherings of engineers. Messages from men distinguished in science, statesmanship and diplomacy were read. Medals struck by the French Government were presented to the general chairman of the deputation, to the chairman of each' individual delegation and to the president's of the four Founder Societies. The meeting was regarded by engineers as formally ushering in a definite movement for world unity in the engineering profession. J. Vipond Davies, president of the United Engineering Society, acted as toastmaster. The speakers ,were Ambrose Swasey, Charles F. Rand, F. B. Jewett, A. S. Dwight, John R. Freeman, all members of the deputa- tion; and M. Gaston Liebert, French Consul General; Captain G. H. Armstrong, British Counsel General; G. G. Clapperton, representing the British Institution bf Electrical Engineers, and Prof. Jacques Cavalier, spokesman for the French universities. Among the engineers from other cities was Governor. James Hartness of Vermont.

Jan 1, 1921

Victor G. Hills, Cripple Creek, Colo. (communication to the Secretary): In continuation of the subject of the perform ance of "baby machine-drills," presented by Mr. Thane, I sub mit the following records of work done with these drills in the Cripple Creek district of Colorado. Two years ago, the Anchoria-Leland mine began to use baby drills ; and, to make the experiment more complete, began by purchasing two machines, of the same size, but of different make. The little drills were first used for stoping and up raising, in which they achieved at once an unqualified success. Soon after, they began to be used for the smaller branch-drifts and prospecting cross-cuts, while the large two-man drills were still used for the main drifts. Now they have almost entirely superseded the large drills, since it has been found that they will do any ordinary work in the mine at less cost, both for labor and for powder, per ton of rock broken. The large drills, of course, make more rapid progress; but it is only where time is a specially important factor that they are now used. The Portland mine, where, until within a year, only the large machines were used, is now using some small drills for stoping, and also for drifting. Mr. Thane speaks of having two men to operate a baby-drill; but in this district the baby-machine is always worked by one man without help. Again, in Prof. Christy's postscript, the quoted extract from a letter of Mr. L. T. Seymour, the South African engineer, mentions a " 2 1/2-inch Little Giant," as though that were the smallest machine used in that region. These statements leave some doubt as to just what Mr. Thane means by a " baby-drill." To avoid any such uncertainty as to the present communication, I will say that the large or two-man drill in common use at Cripple Creek has a 3 1/8-inch cylinder; while a few mines employ a 3-inch, and a few, in running large tunnels, a 3 1/2-inch drill. What we call a baby-drill here has a

Jan 1, 1900

By Benny J. Nelson

AS a part of the fundamental research program of Aluminum Research Laboratories, some data were obtained on the ternary system Mg-Al-Mn. As very little information on the magnesium corner of this diagram has heretofore been published, it seems desirable to make available the values found for the liquidus and solidus surfaces of this system. Procedure The settling procedure was used for the determination of the liquidus compositions. Metallo-graphic examination of quenched samples, and stress-rupture upon incipient melting, were used for the solidus determinations. The settling procedure has been described in a previous paper.' Briefly, this method involved saturating the alloy with manganese at a temperature substantially above that at which the sohbility was to be determined, then cooling the melt to the latter temperature, and holding it at that temperature for a substantial period of time. Samples for analysis .were carefully ladled from the upper portion of the melt at hourly intervals during the holding period. After the ladling of each sample, the melt was stirred to redistribute some of the manganese that had already settled, because it appeared that when the latter particles of manganese again settled, they aided in carrying down more of the manganese and thus hastened the attainment of equilibrium. The melts were prepared and held in a No. 8 Tercod crucible holding approximately 4 lb of metal. The manganese was added either in the form of a prealloyed ingot (Dow M) containing about 1.5 pct Mn or by the use of a flux (Dow 250) containing manganese chloride. In calculating the flux additions, it was assumed that the manganese introduced would be equal to 22 pct of the total weight of the flux. Temperatures were measured with an iron-constantan thermocouple enclosed in a seamless steel tube, the lower end of which was welded shut. This protection tube also served as a stirring rod. The samples ladled from the upper portions of the melts at the various intervals were analyzed for aluminum, manganese, and iron. When making the alloys which were to be used for the determination of the solidus, 2½ in. diam tilt mold ingots were cast, scalped to 2.0 in. in diam, and extruded into ? in. diam wire. The principal impurities in the melts for this investigation were iron and silicon; their total not exceeding 0.03 pct. Portions of the wire, approximately 2 in. in length, were enclosed in stainless steel capsules for protection from the atmosphere. Bundles of these capsules, with a dummy capsule containing an iron-constantan thermocouple, were heated inside a large steel block (acting as a heat reservoir) in a closed circulating-air type electric furnace. At ap- propriate times, the capsules were removed and quenched in water. The wires were examined metallographically to determine the temperature of initial melting. Short times at temperature were used at the beginning for wire specimens of all alloys to obtain quickly the approximate temperatures at which melting could be first observed. When approximate solidus temperatures had thus been determined, equilibrium heating was attempted. This equilibrium heating consisted of an 8 or 16 hr period at a temperature, about 50 °F below the lowest temperature at which melting occurred when short heating cycles were used, followed by further heating for 1 hr periods at consecutive 10" higher temperatures. The theory for the method of stress-rupture at incipient melting has been well covereda and its limitations are recognized. Thus, if the interfacial tensions are such that the first minute quantity of liquid is "bunched up" at the grain boundary junctions instead of spreading out along the grain boundaries,³ temperatures higher than the solidus are required before melting will be manifested by rupture of the specimen. This point will be elaborated later. Specimens of the wires with a reduced section (approximately 1/16 in. diam) were suspended vertically in a tubular furnace. The setup used is shown in Fig. 1. The clamp holding the specimen was made from alumel thermocouple wire and the thermocouple was thus completed across the specimen by attaching a chrome1 wire to its lower end. Temperatures were read from a Speedomax recorder used in conjunction with a calibrated thermocouple. The small weight attached to the specimen and a vibrator attached to the furnace tube, to aid in distributing the molten constituent along the grain boundaries, were used to bring about rupture at a temperature closely approximating the solidus. The specimens were heated at a rate of about 5°C per min. The rupture of the specimens was indicated both by sound and by the action of the recorder. An argon atmosphere containing a small amount of SO² was used for protection of the specimen. The assembly was taken out of the furnace immediately following rupture and the specimen removed. Some of the broken specimens were examined metallographically and will be referred to later. Results and Discussion Fig. 2 shows a set of typical time-composition curves for liquid samples of the Mg-Al-Mn alloys used for the settling tests. The data as presented

Jan 1, 1952

By P. Gianella Vincent

Barite deposits are of widespread occurrence in Nevada but there are few pro-ducing properties; most of the latter are in northern Nevada. The production of the state is small at present-in the neighborhood of 15,000 tons annually-but is steadily increasing. The barite occurs in veins and also as replacement deposits in lime-stone. The bulk of the production has come from the replacement deposits. It is thought that the barite was deposited from rising hot waters similar to-and probably contemporaneous with-those responsible for some of the metallic ores of the state. This paper discusses only the deposits in northern Nevada. THE commercial mining of barite did not begin in Nevada until 1916 but although the production has been small there has been a rapid increase in tonnage during the past several years. Barite is widely distributed in the state and production comes from at least nine of the seventeen counties. Apparently this is not generally known, as only two counties are mentioned in a recent publications dealing with barite. The increase in output is due largely to recently discovered barite deposits in northern Nevada. The principal production has come from the northern portions of Lander and Eureka counties; smaller amounts from Elko County. The discussion in this paper is confined largely to the deposits in these counties. Most of the producing area is tributary to the Western Pacific and Southern Pacific Railroads, whose tracks lie along the Humboldt River. Barite is known from other northern counties but so far the production has been small. The output has come entirely from veins and replacement deposits, the latter largely in sedimentary rocks. In general, the barite from veins is white; that from the replacement deposits usually is dark gray to black, although some is light gray or white. Most of the veins contain sulphides and many barite-bearing veins have been worked for their content of base or precious metals.

Jan 1, 1940

By William Catron

BECAUSE the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for eco-nomic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It was clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stopes below with disastrous results to men and mine. The method of mining developed during the years 1928, 1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled clown gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample to warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost.

Jan 1, 1938

By C. E. Armantrout, J. A. Rowland, D. F. Walsh

THE close-packed-hexagonal structure of mag-J- nesium is converted to a ductile and malleable body-centered-cubic lattice by the addition of lithium in excess of 10 pct. Further, the density of magnesium or magnesium-base alloys is decreased by additions of lithium. The practical possibilities of such alloys as a basis for uniquely light, malleable, and ductile structural materials were pointed out by Dean in 1944' and by Hume-Rothery in 1945.2 It was apparent to these investigators, however, that more complex compositions would be required if strengths sufficient for structural applications were to be developed in these alloys. In a search for strengthening additions, various investigators w have examined a number of the ternary and more complex alloys containing magnesium and lithium. An investigation of the fundamental characteristics of these alloys was undertaken by the Bureau of Mines. The investigation was initiated with a study of the magnesium-rich corner of the equilibrium diagram for the ternary system, Mg-Li-Al. The following data from published investigations of Mg-Li-A1 alloys were available: 1—a description of isothermal sections at 20" and 400°C through the Mg-Li-A1 constitution diagram by F. I. Shamrai;' 2—a diagram by P. D. Frost et al." showing approximate phase relationships at 700°F for a number of the Mg-Li-A1 alloys; and 3—diagrams showing the constitution at 500" and 700°F for the Mg-Li-A1 alloy system published by A. Jones et al.' Where compositions and temperatures permit comparison, these diagrams show disagreement. The 700°F isotherms of Frost and Jones differ only in the placement of the phase boundaries. But Sham-rai's 400°C (752°F) isotherm shows a variation in phases as well as in phase boundaries. Although rigid comparison of these different isothermal sections might not be justifiable, it seems impossible to reconcile Shamrai's construction with the isotherms of Frost or Jones. The isothermal sections presented in this paper were prepared to determine compositions which might be suitable for age hardening and to develop the general slope and placement of the various phase boundaries in the magnesium-rich corner of the diagram. Sections at 375", 200°, and 100°C were selected for investigation. In constructing these sections, the solubility of aluminum in magnesium, as reported by W. L. Fink and L. A. Willey Vn 1948, was used at the binary Mg-A1 boundary and the solubility of lithium in magnesium was obtained from the equilibrium diagram for that system as reported by G. F. Sager and B. J. Nelson" in the same year. The solubility of magnesium in lithium was determined experimentally and conforms in general to data reported by P. Saldau and F. Shamrai." Parameters for AlLi and MgI7A1, were taken from American Society for Testing Materials X-ray diffraction data cards. Experimental Procedures Although the isothermal sections presented in this paper are not unusually complex, the experimental techniques involved in their construction are made extremely difficult by the relatively high vapor pressure of lithium and the great chemical activity of both magnesium and lithium. Because of these characteristics, which make precise control of the composition of equilibrium-treated filings practically impossible, the disappearing phase method was used in preference to the parametric method in conjunction with metallographic studies. The alloys used in this investigation were melted and cast in an atmosphere of helium using a tilting-type furnace which enclosed a steel crucible and mold in a single unit. Each portion of the charge (500 to 600 g) was cleaned carefully just before placing it in the crucible; and the charge, crucible, and entire melting apparatus were evacuated and then washed with grade A helium while preheating to approximately 100°C. The alloys were melted and chill cast in an atmosphere of helium. Alloys prepared in this way were relatively free from inclusions and a fluxing treatment was considered unnecessary. The cylindrical ingots obtained were scalped and then reduced 96 pct in area by direct extrusion, yielding % in. diam rod. Sections of the rod, approximately 3 in. long, were given equilibrium heat treatments and then sampled for metallographic examination, X-ray diffraction study, and chemical analysis. The surface of each equilibrium-treated rod was machined to a depth sufficient to insure removal of contaminated material before samples for chemical analysis or X-ray diffraction study were obtained, and all decisions on microstructure were based on the examination of the central portion of the metallographic specimen. All specimens homogenized at 375°C were analyzed after this equilibrium heat treatment. When the composition of an alloy placed it in a critical area of the 200" or 100°C isothermal section, a check chemical analysis was made on a sample taken from the alloy specimen as-heat-treated at the particular temperature. Standard chemical procedures of gravimetric analysis were used in the determination of magnesium and aluminum; lithium, potassium, and sodium were determined by flame photometer methods

Jan 1, 1956

H. H. Kellogg (Columbia University)—The accurate measurements of the equilibrium gas ratios for the reaction: Pb(1) + H2S = PbS(c) + H2 [I] reported in this paper are a distinct contribution to the literature, and the authors are to be congratulated. I wish to comment, however, on the correlation of the data which is presented. The authors show a straight line correlation of log H2S/H2 vs 1/T, Fig. 1, and are satisfied with this for calculation of ?Fo even when extrapolated to low temperature (427 °C). Heat capacity data for the constituents of this reaction are available (at least to 900°K), however, and ACP for reaction [I] is finite and positive in the temperature range studied. Simplification of the heat capacity equations yields: ?CP= 2.71 + 1.65 x 10-3T (298°-900°K) This positive value of ?CP should result in a small but noticeable curvature of a plot of log K vs 1/T. That the experimental data fail to show such a curvature may be traced to another oversimplification of the author's treatment of their data. They assume that the activity of liquid lead is unity for all temperatures, so that

Jan 1, 1961

By F. L. Ransome

The following paper aims to present in a much condensed form the salient results of a detailed geological study of the Bisbee quadrangle, Arizona, carried on during the latter part of the year 1902. It is essentially an abstract of the final official report which will shortly appear in the series of professional papers of the U. S. Geological Survey, accompanied by maps, sections and illustrations. I. Geography. The Bisbee quadrangle, which includes the Warren mining district, lies in Cochise county in the southeastern part of Arizona. It is inclosed between the meridians 109" 45' and 110" 00' and the parallels 31" 30' and 31" 201, the latter being locally the Mexican boundarg line. The area of the quadrangle is about 170 square miles, and includes the southeastern half of the Mule Mountains, one of the smaller of the isolated ranges characteristic of this part of Arizona. The Mule Mountains, while less markedly linear than the Dragoon, Huachuca, Chiricahua, and other neighboring ranges, have a general NW-SE. trend. They may be considered as extending from the old mining town of Tombstone to the Mexican border, a distance of about 30 miles. On the northwest they are separated by the broad, flat floor of Sulphur Spring Valley from the Chiricahua Range, and on the southeast by the similar broad valley of the Rio San Pedro from the Huachuca Range. The maximum width of the range is about 12 miles, and its greatest elevation attained by Mt. Ballard, west of Bisbee, is 7,400 ft. above sea-level. An index-map, showing the geographical position of the Bisbee quadrangle, is shown in Fig. 1. The town of Bisbee (Fig. 2), with an estimated population of 6,000, is crowded into a few narrow confluent ravines near the

Jan 1, 1904

The plans for the activities of the Petroleum Division for the coming year do not differ materially from those of the past several years. The fall meeting is scheduled for Oct. 12 and 13 and is to be held at Tulsa, Oklahoma. The Division, of course, will take part in the annual meeting of the Institute at New York. A fall meeting will also be held in California at approximately the same time as the meeting at Tulsa. A number of papers are to be presented at the two meetings simultaneously. The various committees are at work arranging the programs. The Division will continue to be active in its various fields of endeavor, such as stabilization, production reviews, and economics, but special stress is to be laid on the engineering and research fields. In the past the Division has served as a forum for open and frank discussion of problems confronting the engineers in the oil industry. It is hoped that it will continue to function in this capacity. Papers presenting original ideas and extending the boundaries of knowledge and understanding in the field of petroleum production will be welcomed. Members who have interesting information to contribute should not wait until called upon by one of the Committees, but should offer a paper at the earliest opportunity. The past year has seen a decided strengthening of the control over production, with prospects of even more rigid control in the future. It is essential that production programs under such control be carefully planned and it is highly desirable that petroleum engineers take an active part in the planning, to insure efficient production of oil and minimize the waste of this natural resource. At a time like this, when it is difficult to keep the production of oil within reasonable limits, it may seem unnecessary to improve the efficiency of production methods, but, as has been aptly said, the trouble is not that we have too great a reserve of oil but that we are producing tomorrow's oil today. Therefore it behooves those who are entrusted with the engineering and scientific phases of production to learn how to get the greatest yield economically possible from the present and future fields in order that we may enjoy the advantages of petroleum products for a long time to come.

Jan 1, 1934

Jan 1, 1961

(Separates of all the Pamphlets published within the last three years (starting with No. 1469) are available, with few exceptions, at Institute headquarters. Separates of all the Technical Publications are also available. All the papers are on file in public, university and technical libraries, and when so indicated in the list, may be found either in the TRANSACTIONS or in the special volumes issued by the Institute.] METAL MINING TRANS. PAMPHLET VOLUME NUMBER TITLE AUTHOR MEETING AND PAGE'S 1033 Relation of Air Pressure to Drilling H. W. Seamon New York 66, 743, 756 Speeds of Hammer Drills Feb., 1921 1034 Time Factor in Depletion of Mines J. W. Roberts New York 66,711 Feb., 1921 1035 Calculation of Ore Tonnage and Grade J. E. Harding New York 66, 117, 122 from Drill-hole Samples Feb., 1921 1038 Discussion of Papers Presented in 1920 1041 Run-off and Mine Drainage H. N. Eavenson New York 66, 624,630 Feb., 1921 1043 The Health of the Underground Worker A. J. Lanza New York Feb., 1021 1045 Dust-ventilation Studies in Metal D. Harrington New York 66, 272, 289Mines Feb., 1921 1046 Discussion of Papers Presented at New York Meeting, 1021 1072 Breakage and Heat Treatment of Rock- B. F. Tillson New York 66,1758 drill Steel Feb., 1921 1073 Discussion of Papers Presented at New York Meeting, 1921 1076-M Analysis of Some Drill-steel Testa F. B. Foley New York 66, 789 Feb., 1921 1077-M Application of Magnetic. Analysis to C. W. Burrows New York 66, 796 Rock Drills Feb. 1921 1078-M Heat Treatment of Rock-drill Steel G.H. Gilman New York 66, 779_ Feb., 1921 1080 What is the Ideal Drill Steel? F. H. Kingdon New York 66,807 Feb., 1921 1051 Rock Drill Steels, etc. - W. T. Ober New York 66,814 Feb., 1921 1082-C Underground Mine Development, Its J. B. Dilworth Wilkes-Barre 66,715, 7.21 Definition and Valuation Sept., 1921 1084 Drill Steel from Hollow Ingots P. A. E. Armstrong New York, 66, 809 Feb., 1921 1088 Authors' Replies to Recent Discussion 1092 Heat Treatment of Rock Drill Steel A. E. Perkins New York 66,811 Feb., 1921 1096-M Octagonal Ventilation Shaft. of Davis- J. L. Bruce Wilkes-Barre 66, 252, 268 Daly Copper Co. Sept., 1921 1100-C Determination of Electrical Equip- G. Bright Wilkes-Barre 66, 580, 605 ment for a Mine Hoist Sept. 1921

Jan 1, 1927

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By A. E. Chute

The shortage of sulfur is not only continuing but appears to be worsening, attended by steadily rising prices. At the same time emphasis on air-pollution abatement is also increasing. These two factors, a rising market and the growth of environmental attention, have encouraged producers to carefully review all possible sulfur sources. As a result, production of sulfur from hydrogen sulfide occurring in petroleum is receiving greater attention than ever before. Until about a decade ago, the practice of recovering sulfur from hydrogen sulfide separated from gases (either found naturally or generated in desulfurization of liquid petroleum) was wide1y recognized but contributed little to total free-world sulfur production. However, during the past ten years, the production of sulfur from this source has increased so rapidly that it now accounts for a substantial portion of total production. This trend shows no signs of slackening, and current forecasts predict that recovered sulfur may become the dominant source of supply within ten years, exceeding mined sulfur production. This paper is intended to list sulfur sources and relationships, and to discuss the potential of producing sulfur from petroleum. The processing involved, in addition to product storage and transportation will also be discussed along with important economic aspects of sulfur recovered from hydrogen sulfide. For many years, most of the world's elemental sulfur has been produced from salt domes, predominantly by the Frasch process. During the past decade, however, another production source has developed so quickly that it is expected to surpass the Frasch production sources within a short time. This new source lies in the petroleum gases and liquids from which dramatically increasing quantities of elemental sulfur are being recovered. In the phraseology of the trade, the sulfur so obtained is know as "recovered sulfur," a term which does not, as one might expect, include sulfur recovered from other sources such as smelter gases. The growth rate of recovered sulfur compared to that of Frasch sulfur, total elemental sulfur, and total Free World sulfur in all forms is shown in Fig. 1 Since 1963, sulfur consumption has exceeded production,with the natural result that the price of sulfurhas continually increased. The posted price reported for f.o.b. Port Sulphur, La., has risen from as low as $18 per ton in 1962 to $32.50 per ton in mid-1967. With freight and other charges, delivered prices have of course run much higher, amounting to $45 or $46 per ton on the east and west coasts. In some cases, spot sales for as much as $60 have been reported. Although it is playing an ever-increasing role in the sulfur market, recovered sulfur was originally a by-product of the removal of hydrogen sulfide from natural and petroleum refinery gases. Today, however, it is often the most valuable product of a multi-product processing facility. Since there are two major sources of recovered sulfur—production and refining processes —natural gas is included with liquid petroleum for purposes of this discussion. Sulfur from these sources is produced in the form of gaseous hydrogen sulfide which is removed from a gas stream and converted to elemental sulfur. Gases in which hydrogen sulfide occurs naturally are natural gas and the gas produced from a high=pres-sure crude=oil or condensate field.

Jan 1, 1969

By D. L. King, C. D. Rubert, R. W. Christy, B. A. Schepman

R. W. Christy (Manager of Sales, Sanitary Engineering Equipment, Link-Belt Co.) and C.D. Rubert (Barrett, Haentjens & Co.) — Messrs. King and Schepman have presented an excellent paper which outlines the general practice in the design of thickener mechanisms. Lifting devices, controls, special applications and underflow pumping arrangements have been reviewed in detail. Pumping arrangements were illustrated which eliminate the need for an expensive tunnel beneath the thickener floor. Most of those illustrated have the common feature of a relatively long underflow suction line extending from the hopper floor to the pump located on the bridge at the periphery of the tank. The authors limit these arrangements to comparatively dilute or thin slurrys. It is implied that dense and coarse concentrates, which are difficult to pump, cannot be handled unless a tunnel is constructed. Fig. 1 shows an underflow pump arrangement which negates the need for a tunnel but which permits the operator to convey a heavy dense slurry. The pump is mounted directly on the fixed hoist platform and the impeller is located at the base of the steel center column. Priming water is not needed. The column

Jan 1, 1963

By M. E. Lombardi

THE Coalinga oil field is located on the west side of the San Joaquin Valley, California. The structure is in general a monocline, the edges of the oil horizon resting on the foot hills and dipping gently toward the east. One prominent anticline occurs plunging southeast. The earlier drilling was done in the foot hills comparatively near the outcrop and the wells were shallow. The sands were followed eastward and, in the case of the anticline, along the plunge, the wells becoming deeper and deeper until the depth of 4,000 ft. was reached and passed. There is nothing to show that the oil will not be found in quantity at still greater depth. In fact, some of the best producers have tapped the sand at close to 4,000 ft. The recovery of oil still farther to the east, and therefore at greater depth, seems to be mainly a question of drilling. In this territory the formations drilled through are chiefly sands and shales; they will not "stand up" in an open drilling hole; the casing has to be carried close to the bit, and it is always difficult to keep the casing free for any considerable distance. Ability to carry casing of comparatively large diameter without conductor pipes for distances of 2,000 or 3,000 ft. or over is desirable in such territory chiefly for two reasons. It makes it possible to enter the oil sand with a pipe of ample diameter; it eliminates one or more expensive strings of casing which act only as conductors for the water string, and furthermore, in territory where waters are encountered which corrode steel rapidly, it makes possible the construction of a rust- and alkali-resisting water string. It is always desirable to shut off top waters, which may lie within 100 ft. or less of the oil sand, with 10-in. pipe. Where the depth is so great that a practical weight of 10-in. pipe will not withstand the probable collapsing pressures, 8 ¼ in. at least is desirable. About the limit of rotary drilling to date in California seems to be the setting of the 10-in. string at 3,200 ft., although the rapid advance in rotary work during the past year seems to indicate that this depth may soon be increased. It is my purpose now, however, to treat only of cable-tool drilling.

Jan 2, 1915