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By W. E. Keck, Paul Jasberg

TheRe is a considerable tonnage of iron ore in the Menominee Range of Michigan that is unsalable only because it has too large a content of sulphur. Beneficiation of such ore is economically desirable, as a relatively large quantity of finished concentrate would be obtained (deleterious sulphur mineral content being generally less than 5 per cent) and because large blocks of such ore have been developed incidentally during search for low-sulphur ore. Research on the flotation beneficiation of these high-sulphur ores consisted of: (1) a study of the flotative properties of the principal mincrals in a nearly pure state, (2) flotation of synthetic mixtures of these pure minerals and (3) application of this experimental information to samples of the ores. Gypsum in iron ores is extremely deleterious; but when sufficiently free from impurities it is a valuable commodity. In 1934, the United States produced 1,500,000 tons of calcined and uncalcined gypsum valued at $13,700,0001. Of this quantity and value, Michigan produced approximately 19 per cent1. Thus this investigation may also prove to be of interest in connection with the gypsum industry if minerals separation, such as flotation, becomes necessary in the future processing of gypsum. Data on the flotative properties of gypsum would also be of interest to the science of flotation because such information has been extremely limited. The technical literature reveals one reference to the flotation of gypsum, in which was discussed the use of the mineral, with several others, to determine the relation between the floatability of minerals and the pH of the flotation solution2. This investigation necessarily was not comprehensive in regard to gypsum, but it did show that the mineral could be floated with oleic acid under varying pH conditions.

Jan 1, 1938

By L. K. Borrett ond C. S. Yusi

The authors have made extensive use of a modern field of mathematics, topology, to give a new approach to the study of sintering. They present a topological model for the sinter body and the void space, and establish a relationship between them. However, in an interdisciplinary paper of this nature it is important that the material be presented, insofar as possible, in

Jan 1, 1967

By John J. Croston

The year 1935 witnessed one of the world's oldest industries—gold mining—attain the stature of a billion dollar business. Preliminary estimates indicate that somewhat over 30,000,000 oz. of gold was produced, the largest output ever recorded and exceeding the previous year by more than 11 per cent. Of this huge total the Witwatersrand field produced about 10,800,000 oz., compared with 10,480,000 oz. in 1934. Russia was apparently the stellar performer, and from incomplete data it appears that between five and five and a half million ounces were produced by the Soviets compared with 4,200,000 last year and 2,814,000 in 1933. The United States nosed Canada out of third place with a production of 3,546,000 oz. against 2,916,000 in 1934, while Canada increased her total from 2,965,000 to 3,291,000 oz. Similarly, gains of lesser magnitude were made in Australia, the Gold Coast and other countries. As a result of great activity in developing outlying areas on the Rand, and the great number of new companies building mills in Canada, Australia, the Gold Coast and the United States, the years 1936 and 1937 should see new high records, as these new producers will more than offset any decline in the older properties. The whole world has become "gold conscious" during the past three years. It has been the subject of innumerable articles by economists and others, but, unfortunately, because of their lack of actual mining knowledge, they have treated the production of gold as a static affair, ignoring geological and technical factors. It is hardly necessary to draw attention to the fact that the world's gold-mining industry was slowly but steadily shrinking in importance, owing to the gradual depletion of known deposits that were profitable to work at costs existing prior to 1929. The International Geological Congress at Pretoria and the League of Nations Committee drew attention to the estimates of the mining experts of the various governments predicting a steady falling off in world production. On the Rand there

Jan 1, 1937

By W. Petruk, R. S. Boorman, R. Gilders

Approximately 33% of the silver in the Zn-Pb-Cu-Ag ore of Brunswick Mining and Smelting Corporation Limited is lost to the tailings and an additional 11% to the zinc concentrate for which no payment is received. 38% to 44% of this lost silver is carried by tetrahedrite, 28% by galena and 28% to 34% by a third silver carrier. The silver minerals acanthite and pyrargyrite have been identified in mill products and hence these are suggested as the principal third silver carrier(s) accounting for 13 to 20% of the silver in Brunswick ore. About 24% of this third silver carrier assemblage is lost to the zinc concentrate and 52% to concentrator tailings.

Jan 1, 1984

By Reuben B. Ellestad, Fremont F. Clarke

In the early days of the lithium industry most of the production was from lepidolite, zinnwaldite, and amblygonite. Nearly all the early extraction processes described in the literature involve heating the finely ground mineral with sulphuric or hydrochloric acid. On subsequent water leaching most of the bases present in the mineral (especially aluminum) are dissolved as sulphates. As a result, the leach solution required extensive chemical purification before the lithium could be precipitated as carbonate. Following the remarkable growth of the lithium industry to its present size, zinnwaldite and amblygonite ores must be considered of minor importance only. Attention is now focused on spodumene, abundant enough in North America to be a major source of supply, and there are important supplies of lepidolite and petalite in Africa. The extraction processes described below all apply to spodumene, although several will also operate on other lithium minerals, such as petalite. Base Exchange with Alkali Sulphates: A distinct advance was made with the disclosures of Wadman and von Girsewalt. In these methods the finely ground silicate ore (spodumene or lepidolite) is intimately mixed with an excess of alkali sulphate (usually K2S04) in at least a 1 to 1 proportion, and the mixture was heated to a relatively high temperature. Base exchange results, with the formation of lithium sulphate. A water leach dissolves the lithium sulphate, together with the excess potassium sulphate. Successful operation of this type of process requires very thorough grinding and mixing, as well as careful temperature control. The use of K,SO, is objectionable from cost considerations since purification of lithium carbonate re- quires the use of potassium carbonate, if the K2S04 is to be recovered and recycled. The lower solubility of K,SO, as compared with Na,SO, is also objectionable, since it limits the concentration of the Li,SO, solution to be precipitated by K,CO,. Early laboratory-scale investigation of this process by Lithium Corp. was not encouraging

Nov 1, 1955

By B. O. Pickard

With more than a score of shafts and numerous stope openings to the surface, all inter-connected underground; with underground temperatures high, often exceeding 100' wet bulb; with an ore presenting a dust health-hazard; with certain rocks giving off inert gases; with an altitude of 6000 ft. above the sea level and a dry desert climate, the Tonopah mines offered an interesting problem to the student of metal-line ventilation. Tonopah is not an old mining district, relatively speaking. It was discovered in 1900 and Tonopah soon was a boom camp. Ry 1910 the mines of this district had paid many millions of dollars in dividends. At the time of the 1921 ventilation study, mining had passed its peak of production and the operators were striving to reduce costs, to discover new orebodies, and to attract desirable workmen. In 1925, the camp was operating on a still less prosperous basis, but the several managements had not relaxed in their efforts to keep the mines on a paying basis, neither had they lost their faith in the ability of the mines and the price of the silver to "come back." Early in the fall of 1921 the Bureau of Mines was invited by the mine operators at Tonopah, Nev., to make a detailed study of the underground air conditions in the Tonopah silver district. D. Harrington, chief engineer of the Safety Service Division of the Bureau, who was then directly supervising the ventilation studies, detailed district engineer B. 0. Pickard, mining engineer E. D. Gardner, surgeon C. E. Kindall, and petrographer IT. Insley to carry on the study. G. S. Rice, as chief mining engineer of the Bureau of Mines, was in general charge of the work. Upon the completion of the study each operator was furnished a report on conditions in his mine. In 1925, a supplemental study at Tonopah was made by the district engineer at the request of several of the operators, and each operator advised on the changes in his respective air conditions and recommendations for improvements. Recently, several of the operators gave the

Jan 1, 1927

By Robert Tally

UNDERGROUND fires have been common in the mines of the United Verde Copper Co. for the past 22 years. The first fire started in the 300 Hampton stope in the fall of 1894, following a cave in that orebody. The soft and highly pyritic nature of the ores is responsible for most of these fires. Since ore characteristics, methods of mining and ventilation, have an important influence on mine fires, they will be described in this paper. Orebodies The ores are mostly chalcopyrite except on the upper levels where considerable chalcocite, black oxides and other secondary minerals occur. The orebodies are in the form of lenses and vary in size. Some are small, while others are several hundred feet in length by a hundred or more feet in width. These orebodies extend, in some cases, from the surface to the lowest levels of the mine and usually at a steep angle of inclination. There are three different classes of ores, classified in accordance with the gangue rock. The copper in all cases, except as above specified, is in the form of chalcopyrite. The largest and most important of these reserves are in a pyrite gangue. Next in order of importance are the schist ores and lastly the quartz porphyries. The pyrite, locally termed iron ores, average from 30 to 42 per cent. sulphur. The schist and quartz, porphyries vary from 3 to 25 per cent. sulphur. Mining Methods All the orebodies were worked by the overhand square-set method of mining prior to 1908; since then the cut-and-fill, shrinkage and mill-hole systems have been largely employed. But 15 per cent. of the total tonnage now extracted is from timbered stopes. Most off this is from the upper levels where the ore is soft, heavy and more or less broken to the surface. Large reserves of these ores are being conserved for more efficient methods of mining.

Jan 9, 1916

By J. M. Glass, C. D. Sholtess

We at Hughes Tool Co. are extremely proud of the quality of the hardware and techniques introduced through our efforts in tunnel-machine development and of the ready acceptance of them by manufacturers and contractors throughout the world. All of us who attended the conference and who are represented in the book, we believe, anticipate an increasing need for improved methods of underground excavation. The development of such methods will permit the further exploitation of our underground resources plus providing underground space for transportation and habitation at a rate compatible with demand. We hope that our comments, derived from the application of Hughes Tool Co.'s rotary boring experience, may provide a better understanding of the boring method as a possible avenue through which the desired improvement in underground excavation may be attained. We hope you may also become aware of some of the problems Hughes, as a pioneer developer of raise, shaft, and tunnel-boring equipment, must face prior to further major undertakings in this very interesting field. Our task is to acquaint you with the present status of the Hughes Tool Co. development program as it relates specifically to tunnel-boring machines. We feel this task can best be accomplished by a brief review of the general sequence of events leading to and constituting an equipment development program and more specifically the manner in which these events have occurred to either influence or further our boring-machine development program. The development of a machine to accomplish a certain function or series of functions is usually a very long and tedious process to which any number of individuals must contribute, perhaps over a period of centuries. Of particular irony is the fact that often those who contribute most generously do not necessarily reap the greatest immediate benefit. A marketable ma- chine will, however, result from a sequence of events such as these: 1) A problem or need is defined. 2) An idea is born to eliminate the problem.

Jan 1, 1970

By C. R. Sundeen, C. W. Nicolson, J. D. McAuliffe, C. A. Campbell, P. D. Pearson

SHAFT SINKING SYMPOSIUM is a weighty term that could mean any number of things. However, in, this case it has been possible to utilize the transcription of the Annual Meeting Session of the same title and present it in much the same manner as the experts on the panel presented it to their audience Wednesday afternoon, February 18, 1953, at the Los Angeles Statler Hotel.

Jan 8, 1953

By D. R. Babich, R. A. Allwes, L. V. Wade

The design of a longwall entry system is discussed in this Bureau of Mines paper. This is done primarily through the results of an in-mine case study made of two sets of longwall entries utilizing the stiff- yield pillar design concept. The rock mechanics instrumentation plan used in this case study is shown, including a description of the instrumentation layout and the types of instruments used. The analysis of the stress changes in the abutment and yield pillars is presented in detail as the longwall panels were mined. Conclusions concerning when the pillar first started to experience significant stress, pillar stress history as the longwall face approached and moved past the pillars, stress distribution within the pillars and relative stress changes between yield and abutment pillars are discussed. The paper concludes with a discussion of the redistribution of stress in the vicinity of the long- wall panel.

Jan 1, 1982

By Arthur Hockman, Howard W. Jaffe, Howard F. Carl

THE problem of establishing practical techniques for evaluating natural corundum ores arose from the desire to improve the existing purchase specifications for crystal corundum procured by the Federal Government for stockpiling. Such purchase specifications had been developed for a large number of products, including minerals and ores considered to be critical and strategic because of the country's dependence on foreign sources of supply. It will be apparent to anyone considering the subject of purchase specifications that the determination of whether or not a submittal meets the specifications is a prime consideration. Consequently satisfactory specifications for minerals should not only incorporate physical, chemical, and mineralogical requirements that are practical but should also establish the means by which both buyer and seller can be assured that materials meet stated requirements. The Munition Board's Material Purchase Specification for corundum, dated February 1954, covered two grades of crystal corundum intended primarily for abrasive use. It established limits of size and of composition: Grade A, at least 95 pct +3 mesh and 92 pct alumina (Al2O3); Grade B, at least 95 pct +7 mesh and 90 pct alumina. It also specified that "of each lot of material, not less than 95 percent shall have a hardness equal to or exceeding nine (9) on Moh's Scale." These specifications were unsatisfactory for several reasons. They did not allow the purchase of lower grades of corundum being used by industry. Neither did they specify exactly what a corundum ore actually was, except that it should contain so much alumina and have a certain scratch hardness. Nothing was said about the abrasive qualities of the material that was to be purchased primarily for abrasive use. During the period 1946 to 1948 supplies of the two grades defined in the then current specifications continued to decrease. Early in 1949 industry recommended that the finer sizes of corundum known as "crystal corundum concentrates," ¼-in. to 16-mesh material, be added to the stockpile specifications and that lower grades of crystal corundum also be included. In the spring of 1949 the National Bureau of Standards and the Bureau of Mines were asked by the Munitions Board to cooperate in developing more practical specifications for grading natural corundum ores. The National Bureau of Standards indicated that it would investigate physical test methods for determining abrasive characteristics of corundum, and the Eastern Experiment Station of the Bureau of Mines at College Park, Md., stated that it would consider the mineralogical aspects of the problem. Members of these two agencies, in consultation with representatives of the Office of Materials Resources of the Munitions Board and of the Federal Supply Service of the General Services Administration, and with the cooperation of industrial suppliers and consumers of abrasive corundum, developed, in a fairly short time, more practical specifications for six grades of corundum ores. The new specifications, issued January 1951, were for corundum, natural crystals, as by this time high-purity synthetic corundum was available and widely used. Nevertheless, for certain specific applications the natural product was preferred over the synthetic material. It has been extensively used in forming snagging wheels. It breaks into more or less cubical particles, whereas artificial corundum fractures into splintery shapes. These produce irregular cuts in glass, while the more symmetrical shape produces a more uniform ground surface. For this reason natural corundum has been preferred for certain types of lens grinding. Pure corundum is clear, colorless alpha alumina (A12O3), but it is rarely found in nature in this form. The most important gem varieties include the blue and yellow sapphires and the ruby. Minute impurities of titanic oxide impart a blue color,

Jan 1, 1955

By K. Verghese, R. W. Rousseau, R. P. Gardner

A radioactive tracer method and a size-discretized, mass balance model are formulated and demonstrated in a feasibility experiment for the determination of crystal growth rates in MSMPR crystallizers. The system studied is aqueous KN03 with K-42 as the radioactive tracer. The results indicate that the method should prove to be very valuable for this application, particularly when systems exhibiting size-dependent growth rates are of interest. An indicated refinement to the tracer method is the simultaneous use of an additional tracer in the liquid to determine the amount of radioactivity from the liquid that is adsorbed on the samples of filtered crystals.

Jan 1, 1980

By John J. Croston

THE year 1935 witnessed one of the world's oldest industries-gold mining-attain the stature of a billion dollar business. Preliminary estimates indicate that somewhat over 30,000,000 oz. of gold was pro-duced, the largest output ever recorded and exceeding the previous year by more than 11 per cent. Of this huge total the Witwatersrand field produced about 10,800,000 oz., compared with 10,480,000 oz. in 1934. Russia was apparently the stellar performer, and from incomplete data it appears that between five and five and a half million ounces were produced by the Soviets compared with 4,200,000 last year and 2,814,000 in 1933. The United States nosed Canada out of third place with a production of 3,546,000 oz. against 2,916,000 in 1934, while Canada increased her total from 2,965,000 to 3,291,000 oz. Similarly, gains of lesser magnitude were made in Australia, the Gold Coast and other countries. As a result of great activity in developing outlying areas on the Rand, and the great number of new companies building mills in Canada, Australia, the Gold Coast and the United States, the years 1936 and 1937 should see new high records, as these new producers will more than offset any decline in the older properties. The whole world has become "gold conscious" during the past three years. It has been the subject of innumerable articles by economists and others, but, unfortunately, because of their lack of actual mining knowl-edge, they have treated the production of gold as a static affair, ignoring geological and technical factors. It is hardly necessary to draw attention to the fact that the world's gold-mining industry was slowly but steadily shrinking in importance, owing to the gradual depletion of known deposits that were profitable to work at costs existing prior to 1929. The International Geological Congress at Pretoria and the League of Nations Committee drew atten-tion to the estimates of the mining experts of the various governments predicting a steady falling off in world production. On the Rand there

Jan 1, 1936

By W. J. Sharwood

The assay or estimation of the total gold content of an ore presents little difficulty, when circumstances permit of securing a thoroughly representative sample. The general practice has been fairly standardized, and such errors as occur in the results obtained are for the most part attributable to the methods of sampling rather than to those of assaying, while duplicate assays made by skilled workers usually agree closely. When it is a question of estimating the "free" or amalgamable gold of an ore, as opposed to the total gold, there is but little uniformity in the methods employed by different workers, and often equally little in the results obtained when duplicate determinations are attempted, no satisfactory standard method having been generally adopted. The methods in use are of two general types—"panning" and "amalgamation." Panning For the estimation of the value of the visible gold by panning or washing, a weighed sample may be taken, ranging from an ounce to several pounds; a kilogram or 2 lb. is a favorite quantity for laboratory tests, while a prospector often uses a full miners' pan, say nearly 20 lb. or 10 kg. In the case of rich quartz, ground to about 30 or 40 mesh, a Mexican horn is convenient, using 1 to 8 02;. of material, and a 6-in. frying pan is favored by many prospectors in dry districts. Bateas, plaques, and dishes of various patterns are used for the same purpose. With these, fairly close results may be obtained by using for ocular comparison standards consisting of weighed amounts of gold of similar character as regards purity, fineness of division, and the general shape of the particles, which standards are preferably obtained from gold of the same locality, or the same deposit, as the ore or gravel being tested. Men working in one particular district sometimes become surprisingly expert in thus judging the value of an ore by the eye, the weights of both ore and gold being guessed. On the other hand, many have been led to overestimate the value of ore and gravel from unfamiliar localities, where the gold has been very finely divided or of unusually high silver content, and a novice may

Jan 1, 1916

By Albert F. J. Bordeaux

The following general survey of the character and present condition of the silver-mines of Mexico, though not offered as a detailed treatise, has been prepared with care, is believed to be generally up-to-date, and is offered as a summary which may be found useful by mining engineers. For several centuries Mexico has ranked first anlong the silver-producing countries. It is now second as a copper-producer, and seventh as a gold-producer, with a fast-increasing output every year. The total amount of capital invested in its mining- and smelting-works exceeds $250,000,000, while the value of the yearly metallic production is $160,000,000. There are more than 20,000 recorded mining properties. The silver-mines of Mexico represent deposits of three principal classes: (1) veins and stock-works in eruptive rocks; (2) fissure-veins crossing the crystalline schists, granite, diorite, etc.; and (3) veins and replacement-deposits in limestone. The true fissure-veins in some districts contain (in value) vol. xxr1x.—22

Jan 1, 1909

By S. K. Tarby, W. O. Philbrook

Limited experimental data and a critical review of the literature are given to indicate that the true equilibrium distribution of manganese between carbon-saturated iron and blast-furnace type slags has not yet been satisfactorily established as a function of the composition, temperature, and pressure of the system. THE distribution of manganese between carbon-saturated iron and blast-furnace type slags has been studied by a number of investigators.1-7 Generalization of the results of these works indicates that the weight percent distribution ratio, (%Mno)/[%Mn], decreases with increasing temperature and slag basicity, in accord with thermodynamic data and practical observation. However, where comparisons among these studies can be made for similar slag compositions and temperatures, the reported values of the manganese distribution ratio are discordant. The variations are attributed to failure of the experimenters to bring their systems completely to equilibrium with graphite and CO at 1 atm, as judged by departures from silicon equilibrium. Even in the most recent work,5-7 this problem was recognized but the authors made the questionable assumption that equilibrium could prevail for manganese distribution even though other reactions were still in progress. Manganese equilibrium values for three particular slag compositions were required incidental to a kinetic study8 and were determined experimentally. Special pains were taken to approach concurrent equilibrium with respect to silicon. Although the results were not completely definitive, they indicate that the true manganese equilibrium distribution ratios are lower than have heretofore been reported. The data are presented together with a brief commentary on the literature to suggest the need for further work in this area. EXPERIMENTAL METHOD Equilibrium runs were carried out in a stationary, induction-heated graphite crucible closed by a graphite lid and flushed continuously by a slow flow of carbon monoxide at 1 atm pressure. Temperature was controlled manually within ±3 deg C with the aid of immersed thermocouples in mullite tubes protected against slag attack by graphite sleeves. The desired temperature was established by a calibrated Pt vs Pt-10 pct Rh thermocouple and control was then maintained using a more stable W-Mo couple. Runs were made at 1500º and 1575ºC for durations of 7.5 to 10.5 hr. The charges were calculated to yield 400 g of metal and 100 g of slag. The primary slag compositions and designation numbers were as follows:

Jan 1, 1963

By D. A. Hinckfuss

The large and compact but simple Bougainville Copper Limited flotation circuit has been in operation for 3-1/2 years. Discussion of problem areas encountered since start-up and remedies that have been implemented or proposed, form the main basis of this paper. Full current operational data is also provided.

Jan 1, 1976

By Carl Floe

THE possibility of separating the various constituents of mixed ores or metalliferous products by differentially producing their soluble sulfates has not received the consideration that it deserves. The production of the soluble sulfate may be accomplished by baking the ore with sulfuric acid or by treating it with gases containing oxides of sulfur. This paper is primarily concerned with the first method. The procedure used was to mix the material with a solution containing a definite weight of sulfuric acid, baking the mixture at various tempera-tures and then leaching out the soluble sulfate with water. The amount of acid used was calculated to produce certain desired soluble sulfates. The temperature of baking depended on the decomposition points of the sulfates involved. The object was to prevent the formation of or to decompose certain sulfates while leaving a maximum quantity of another sulfate in a water-soluble form. This procedure has two possible advantages over the ordinary method of leaching with an aqueous solution of sulfuric acid. First, a more selective leach may be obtained, and second, the higher temperatures and stronger acid concentrations involved will permit a better recovery of materials that are difficultly soluble in relatively dilute aqueous sulfuric acid solutions. The method of sulfuric acid baking followed by water leaching is not new, as it has been applied in the treatment of various metalliferous materials such as the recovery of cadmium from flue dust1, the re-treat-ment of zinc leach residues2 and the working up of electrolytic slimes3.

Jan 1, 1936

By Frank E. Shepard

Electric power in mining-operations is now successfully applied to haulage, hoisting, lighting and pumping; and until lately, drilling was the one department of mining in which an electric source of energy had not been introduced. Drilling by electric power has been the field for much investigation and several forms of electric-drills have been tried; but in the severe service of rock-drilling many dieculties have been encountered, and the heavy cost of repairs has caused a deep prejudice against this system of drilling. In the majority of electric rock-drills previously placed on the market, two principal factors have been considered; first, a portable electric motor usually of heavy pattern, coupled to the driving-mechanism of the drill proper by means of a flexible shaft; and second, the coupling of some form of reciprocating cross-head to a piston by means of a spiral spring, the drill-bit being fastened to the piston-rod by some suitable form of chuck with fastening bolts and nuts. In this latter form of drill, the cross-head has a fixed stroke, while the piston with drill-bit attached is allowed a variable stroke through the

Jan 1, 1904

By H. E. Gross

Recently, areas of abnormal rock pressure have been drilled on the Gulf Coast of the United States. These pressures may be caused by the recent and pronounced raising of some of the structures without loss of the formation pressure that existed when the reservoir was developing and the producing sand much more deeply buried. It is possible also that a permeable fissure may connect a shallow accumulation of gas with the gas cap of a much deeper reservoir of oil, gas and water. The pressures are so great that dependence on blowout-prevention equipment alone is hazardous to life and property. Fortunately, wells with excessive casing-head pressures, such as the Stanolind Oates No. 1 of the South Houston field, an 8700-ft. gas well with 6000 Ib. per sq. in., and the near-by Stanolind Candelari No. 1, and 8100-ft. gas well with 5000 Ib. per sq. in., have not blown out and cratered. A case is on record of drill pipe in a well in Harris County, Texas, (Oates No. 1) rotating for 8 hr. through blowout-prevention rams holding a back pressure of 2900 Ib. per sq. in. at the casinghead after the mud was altered by salt-water intrusion. Such conditions warrant the use of extremely heavy muds having low viscosity, made from galena (PbS) or specially prepared mixtures of hematite (Fe²O³) and galena suspended by bentonite. In some Gulf Coast wells, sodium silicate mixed with concentrated brine is used to drill through heaving shale formations. Such a fluid inhibits heaving remarkably, but it is expensive and costs as much as $10 per barrel. Barite has been used to weight this mixture. However, this mineral too rapidly increased an already excessive viscosity and it was thought that galena might be better for weighting such drilling fluid. The operator was anxious to try the latter, but none was available in the district. Laboratory Study of Heavy Muds To foster the use of the plentiful and extremely heavy mineral, galena, research has been done recently at the University of Oklahoma to determine the mud-making properties of this substance. For comparative

Jan 1, 1938