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R. W. Guard (General Electric Co., Schenectady)— The method of fractional replication discussed by the author has proven very valuable to us in applied research connected with alloy and process development where many of the same experimental difficulties exist. Several of our experiences may be of assistance to those who want to apply the method in these fields. Contrary to the author's observation that interactions are rarely significant, we have found in a number of cases that first order interactions exist which are significant and metallurgically important. Because of this difference there are two further precautions which should be taken in order to minimize the difficulties from this point. First, in evaluating the importance of

Jan 1, 1957

Discussion of the paper of Norman B. Braly, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 80, August, 1913, pp. 1881 to 1906. GEORGE A. PACKARD, Butte, Mont.:-I would like to ask Mr. Braly if he has any information as to the relative efficiency of the sinking pump, using it as he refers to it, discharging the exhaust into the water column or otherwise. MR. BRALY :-I have never gathered any information on that point. It is a question of getting the work done. In driving the shaft, using that method, we find it easier on the men and everything else. In the bottom of the shaft, if you exhaust back into your column, it is a quiet. shaft. Another thing is the freezing: if you are sinking with air in your pump, you will find it is liable to freeze, and in that case as the air is still under compression it will not freeze the pump. The difference is very small if there is any. But, as might be expected, the pump is a little less efficient when exhausting the air into the column. MR. PACKARD :-I suppose all of us know what a pump in the bottom of the shaft is. My experience has been, although I have no figures, that the capacity of a pump is quite considerably decreased if you are pumping, say, from a depth of 200 ft. through a No. 7 Knowles pump, when you exhaust the air into a column, but, on the other hand, it is also my impression that we can raise the water to a little greater distance. I would like to get some figures from somebody who has made actual tests.

Jan 11, 1913

By Thomas Bannon

PROSPECTING for oil in Spain has been going on for many years but only in a very haphazard way and without adequate geological or technical advice. Shallow wells have been drilled in several provinces but, for the most part, they have been doomed to failure from the beginning because of unfavorable geological locations, inadequate drilling equipment inexperienced operators, or lack of sufficient capital. Geologists seem to agree that the most favorable chances for commercial fields are in the northern part of Spain where the petroliferous Triassic beds underlie Cretaceous and Tertiary sediments. The prov-vinces of Santander, Burgos, Alava, Vizcaya, Navarra, and Guipuzcoa seem to offer the most promising localities. Several shallow wells have been drilled near Villamartin about 60 Km. northeast of Cadiz. One of these wells produced a small quantity of light oil but was never operated. The Spanish Government has been drilling on a well in this district for some time, using the Russian free-fall drilling system. The present depth is about 3000 meters. Several years ago the Compania Petrolera Franco-Espanola drilled a well to a depth of about 700 meters near Elorrio, province of Vizcaya. A diamond-core drill was used. The well was located on vertical Cretaceous beds near an igneous dike. The same company, in March, 1923, made a new location about 500 meters away from the first well and the same distance from the dike, in beds dipping about 80°, and is now drilling its No. 2 well. The last depth reported was about 200 meters.

Jan 3, 1924

By J. D. H. Strickland, K. J. Jackson

IN previous papers' ' details were given of the constructlon and use of an apparatus to study the rate of chlorine consumption and the rate of sul- fate and sulfur production when dilute aqueous chlorine solutions were allowed to react with pyrite and galena in acid solutions at or near room temperature. The reaction with pyrite was very rapid, and hence entirely transport-controlled in its kinetics, and produced only ferric and sulfate ions. The behavior of galena was much more complex and was studied in some detail. Sulfur monochloride could be detected as a reaction product and both

Jan 1, 1959

By Royal Jackman

Two comprehensive papers have appeared regarding the forms of copper that occur in smelter slags, one by Frank E. Lathe1 and the other by C. G. Maier and G. D. Van Arsdale.2 These authors comment on other published results and draw conclusions from their own investiga-tions. The development of X-ray examination of minerals and improve-ment in high-power microscopic methods make it possible today to get additional evidence by means that were not available to earlier investi-gators. The present work was confined to reverberatory slags obtained from five different plants. Some definite conclusions were reached regarding the forms of copper present in these slags. In most cases these agree with some of the findings of Maier and Van Arsdale but new methods of attack were used in the investigation.

Jan 1, 1933

By M. A. Scheil

Austenitic stainless alloys are susceptible to stress-corrosion cracking which may occur under certain corrosion environments irrespective of their susceptibility to intergranular corrosion. Test samples are described for investigating stress-corrosion susceptibility of alloys for industrial equipment. A laboratory test using a solution of magnesium chloride boiling at 309 F. was found to produce transgranular cracks in several alloys investigated and in types 347 and 316 stainless steel tubing a residual stress of the order of 10,000 psi. was observed to initiate cracking. The susceptibility to stress-corrosion cracking as judged by the boiling magnesium chloride test indicates that alloys may be selected which will withstand a high stress in service Stressed specimens of stainless alloys included in a corrosion testing program will aid in determining the acceptibility of these fabricated alloys, when stressed, to the service conditions.

Jan 1, 1945

By P. E. Phillips

INTRODUCTION Rocky Mountain Energy is the mining subsidiary of the Union Pacific Corporation, and has been active m uranium exploration since 1966. Since 1970, RME's uranium activities have been funded 50% by Mono Power, the fuels supply division of Southern California Edision. RME and Mono recognized the need for development of new methods for recovering low grade uranium deposits, and contacted Halliburton Services in regard to in situ leaching, because of Halliburton oil industry experience. As a result, a joint venture between the three parties was formed in 1974. BEAR CREEK The first ISL test was at the Bear Creek deposit in the Powder River Basin of Wyoming, about 60 miles north of Douglas. The mineralized sand is the tertiary Wasatch formation, which is an arkosic sandstone. The sand typically contains about 10% clay. The predominant clay being montmorillonite with some kaolin and illite. The roll front type trends occur along the flanks of old stream channels. This deposit was being considered for exploitation as an open pit mine. Routine amenability testing indicated it was quite amenable to a conventional acid leach, consuming about 50 lbs. of sulfuric acid per ton of ore in an agitation leach, while recovering more than 95% of the uranium. Hydrology testing indicated a permeability of about 1 darcy. The groundwater has a pH of about 8. 0 and TDS of about 1200. Test Facility The Bear Creek test was a summer-only operation, with no effort towards weather-proofing. Uranium was removed from the recovered leach solution by fixed bed ion-exchange columns. Loaded resin was eluted with strong sulfuric acid. The recovery plant was designed for simplicity of operation, with little concern for the grade or form of the yellowcake product. The plant was constructed of PVC piping and valves, fiberglass reinforced tanks, and stainless steel pumps. All waste streams were directed toward a lined evaporation pond. Wellfield The test pattern was a 30 foot radius 5 spot, with wells fully penetrating the 30 foot ore sand to a total depth of about 165 feet. The injection wells were three-inch PVC with slotted casing, while the producer was eight-inch I. D. , with a stainless steel screen. The eight-inch well was necessary in order to accommodate a six-inch stainless steel pump. Wells were cemented using cement baskets, at the top of the sand, to contain a latex acid resistant cement. Leach Conditions The lixiviant generally contained from five to seven gpl sulfuric acid, and about 0. 5 gpl hydrogen peroxide. The pH was generally 1.4 and EMF was usually -550 millivolts. Injection rates were rather variable from well to well, with a total injection rate between 12 and 25 GPM. Production rate was set at 25 GPM, since the production pump could not be throttled more than that. Approximately two pore volumes of solution were injected during the three-month test. Leach Results Injectivity was quite variable, but was generally rather poor, because of poor completion techniques. The problem stemmed from injecting before the injection wells had been completely cleaned of drilling mud. Pregnant solutions averaged about 50 ppm U308, with groundwater dilution varying from 10 to 50%, depending on the injection rate. Injectivity did decline with time, probably as a result of contamination of the well faces. A total of 1000 lbs U308 had been recovered when the test was shut down. This probably represented about 10% of the mineral under leach.

Jan 1, 1979

By J. D. Prater, H. R. Spedden, E. E. Malouf

With the expansion of copper leaching of the various mine wastes at the Kennecott properties, additional copper recovery facilities have been required. A research and development program has been pursued to determine a suitable economic approach to the processing of the additional leach solutions contemplated. lnvestigations on laboratory and pilot plant scales covered electrowinning, solvent extraction, launder and vat type cells, activated launders, precipitation drums, and precipitation cones. This paper contains information on the development and use of a new precipitation-cone-type copper recovery plant, which provides many advantages over older systems. The application of research findings to the old art of leaching copper from copper-bearing mine waste has resulted in a significant contribution of copper to over-all copper production. For example, at the completion of the current expansion program at the Kennecott properties in the United States, copper produced from waste dump leaching will amount to about 25% of the total production. Formerly only 2% was derived from this source. This planned expansion of copper leaching focused attention on the problem of developing more efficient methods for the recovery of copper from greatly increased volumes of copper-bearing solutions.1 The detailed investigations which have been made in laboratory pilot plant and plant tests for the recovery of copper from copper-bearing solutions have included electrowinning, solvent extraction, ion exchange and cementation with iron in launders, precipitation drums, activated launders, and precipitation cones. The results of these investigations have led to the development of a new precipitation cone-type copper recovery system providing many advantages over older methods. ELECTROWINNING OF COPPER FROM COPPER-BEARING SOLUTIONS The direct electrowinning of copper from relatively dilute solutions of copper-bearing mine water, namely solutions containing 10 to 20 lbs of copper per 1000 gal, has always been an attractive possibility.2,3 By this method, high-purity copper powder can be recovered with a power consumption of 3.5 kwh per Ib of copper. However, once the solution strength de-

Jan 1, 1967

By J. G. Bartram

Jan 1, 1934

By Richard Akerman

At the International Sessions of the Iron and Steel Institute of Great Britain, the American Institute of Mining Engineers and the Verein Deutscher EisenhUttenleute, held in Allegheny City, Pa., in October, 1890, Sir James Kitson, then President of the Iron and Steel Institute, presented an account by Sir Henry Bessemer of the origin of the process which bears his name.? The paper graplhically told how, after the first successes immediately following the first reports given at Cheltenham of the new process, there rapidly succceded such obstacles and difficulties in its practical application, that the whole matter was nearly abandoned in 1858. Suddenly, however, the situation changed, so that at length it became possible

Jan 1, 1894

IN my opinion antimony* is a composition made by Nature to create a metallic mineral that is overflowing with an undue proportion of hot and dry material and with its moisture poorly mixed, with an effect wholly contrary to the composition of metals. Therefore, it comes to be, like quicksilver, a mineral deformity and monstrosity among metals. Or it might be a material that is about to reach metallic perfection, but is hindered from doing so by being mined too soon. I am persuaded to this opinion by seeing in it so many parts similar to those of metals, considering first its clear and brilliant color and its great weight and then the regulus† that it leaves at the bottom of the vessel on melting. Obviously it has in it much earthiness, as is attested by the odor of its sulphurous burning, its imperfect blending, its defective mixture, its difficult fusion, and finally the character of its regulus, for although this is very white and almost more shining than silver, it is much more brittle than glass. [28] The practical philosophers make great use of it in their alchemistic work in order, they say, to make an oil that they believe gives the color of gold to fixed silver. For this reason many of them hold it in high repute, especially the kind that has long thin fibers like a bundle of bristles. And they say that they can obtain greater results with it than with sulphur. I believe this, however, in case it is aided by sulphur, since it is nearer the metallic species than sulphur is. This I know for a certainty because I have seen them extract from it a bloody liquor in the form of an oil. The alchemist who made that which I saw told me that it had not only the

Jan 1, 1942

By Alan Stanley

The MacIntyre Development of the National Lead Co. is located at Tahawus, N. Y. The operations involve the mining and concentrating of a titaniferous iron ore to produce an ilmenite concentrate and a magnetite concentrate. Construction of the MacIntyre plant was commenced during the summer of 1941, when world conditions threatened to cut off the supply of Indian ilmenite. An open pit mining operation was developed and the crushing and milling equipment put in operation in July 1942. A general description of the operation was given in the Adirondack Issue of Mining and Metallurgy for November 1943. The metallurgy of the mill operation was described by Mr. Frank R. Milliken,* Plant Manager, National Lead Company, MacIntyre Development, and presented at the AIME New York Meeting, February 1948.

Jan 1, 1949

THE Copperton mill in reality was a sort of proving ground. It was. designed to serve three purposes: (1) to verify the accuracy of the mine sampling by actually treating substantial tonnages of ore, (2) to demonstrate on a reasonably large scale the percentage of the copper in the ore that could be recovered, and (3) to permit the testing of various kinds of machines and devices for crushing and concentrating the ore, so as to guide the engineers in designing the proposed 6000-ton milling plant at Garfield. As a matter of fact, it contained a lot of second-hand equipment sent by MacNeill and Penrose from the works of the United States Reduction &Refining Co. in Colorado; and it was by no means a model installation. Probably it would have been even less efficient, except for the skillful work of George O. Bradley and-Frank-G, Janney, both of whom had worked with Jackling at Mercur. Bradley was a mechanical engineer employed to design the plant. Later he directed the design of the concentrator at Garfield and remained as consulting mechanical engineer to the company until 1915. Janney, who afterwards became manager of mills, a position that he retained until his death in 1916, was a skilled mechanic at the Golden Gate mill at Mercur when Jackling built the plant in 1897. His flair was making machinery do what it was intended to do. He was not an ore-dressing engineer in a technical sense; but Jackling him- self was a specialist in this respect as well as in designing and construction and Janney's mechanical aptitude made him a particularly valuable man in the Utah organization. John McDonald, another of Jackling's Mercur associates, was the first mine superintendent. With practically all of the available funds earmarked for construction of the Copperton

Jan 1, 1933

By T. R. Scott

The acid pressure leaching of zinc-lead flotation concentrates from the McArthur River deposit, Northern Territory, Australia, has been studied in a three-stage, continuous, co-current system, using titanium reactors of 6 gal capacity. The starting pulp contained 2.7 lb solids per gal of synthetic spent electrolyte and the final liquors contained 120-125 gpl Zn. Parameters investigated included temperature (95-120°C), oxygen pressure (30-100 psi), acid addition (80-130% of theoretical), and leaching time (35-65 min per stage). Technical factors studied included mixing and pumping of feed, agitation, settling and washing and control mechanisms. The chemistry of liquor impurities, of scale formation in autoclaves and of lead jarosites in residues was studied in some detail. Results indicated that 90-93% zinc recovery could be obtained with a "low- acid" process, using 113°C and 70 psig with oxygen as the favored operating conditions. Residues contained 16-18% elemental sulphur and 15% lead, together with unreacted pyrite and gangue minerals. Pelletizing of residues was impracticable, and alternative suggestions for the recovery of lead, sulphur and silver are discussed. Preliminary estimates for plant and operating costs are given and enable some comparison to be made with conventional procedures.

Jan 1, 1973

By Per Larsson

The Chapin Mine, on the Menominee range, Lake Superior, was first opened in 1880 and has since then produced 1 1/2 million tons of soft blue hematite, containing about 63 per cent. of iron and 0.07 per cent. of phosphorus. Up to the present time, the ore has been mined from three separate lenticular deposits, which conform with the Huronian strata of the district, striking N. 75' W., dipping from 70' to 80' N., and pitching about 30' W. The largest of these ore-bodies has a horizontal length of 2500 feet, and a maximum width of 130 feet in the middle, from which it gradually narrows to the pointed ends. Its dimensions have not changed materially as greater depth has been attained. Figs. 1 and 2 show the geology and the mine-workings on the main deposit.. As this main deposit pitches away from the eastern part of the property, it is replaced by another lens, containing ore of the same quality. When first opened on the 1st level, this lens was 150 feet long and up to 50 feet wide, while now on the 6th level it shows a length of 550 feet, and a maximum width of 100 feet. It is separated from the extreme east end of the main deposit by about 30 feet of clay-slates It pinches out gradually between the slates to the west, and stops abruptly against the jasper to the east. Farther east in this belt of jasper occurs the third ore-lens which has a pretty regular length of 100 feet and averages, from the 1st to the 5th level, 20 feet in width. A cross-section of the ore-formation shows, on the north or hanging side of the ore, about 200 feet of clay-slates, and farther north a heavy belt of magnesian limestone. The slates and the dolomite are generally separated near the surface by a conglomerate of broken dolomite and soft slates. The foot-wall rock is also a clay-slate, containing a higher percentage of iron and less magnesia than the hanging-slate. Farther south occur alternate beds of slates and lean ore or jasper.

Jan 1, 1888

By Fred Bond

WHEN some materials are ground dry in a ball mill, a stage of comminution is reached at which the finely divided particles begin to adhere to the balls and to the mill lining. As grinding progresses, a coating accumulates upon the grinding media, which tends to cushion the impacts and thus retard reduction of the material, and finally becomes so thick that all grinding ceases. At this stage, the characteristic noise produced by the grinding media dur-ing normal mill operation changes to a dull, muffled sound. In most instances, the formation of the coating marks the practical limit of size reduction. Industrial requirements for fine grinding frequently meet with this obstacle, and the tendency in many com-mercial operations to grind materials to a higher degree of fineness than formerly makes the solution of this problem an im-portant one. The phenomenon is of far greater importance, therefore, than the limited amount of research and technical discussion devoted to it implies. In order to learn more about the coating of different materials when ground dry to a high degree of fineness, certain labora-tory tests were made from which it was found that different materials form coat-ings at widely different size distributions. It is probable that most materials will coat if they are ground fine enough; the exceptions have inherent lubricating qual-ities, like coal and graphite.

Jan 1, 1940

Secretary Membership Committee Increase of Membership Papers and Publications Committee Library Committee Treasurer Report of the Secretary TO THE BOARDOF DIRECTORASN D THE MEMBEROS F THE AMERICANIN STITUTE OF MINING AND METALLURGICAENL GINEERS GENTLEME:N As will be shown by the special reports of the Finance, Membership, Library, Papers and Publications, Increase of Membership, and other standing Committees, together with that of the Treasurer, the year 1927 was one in which the Institute made very satisfactory progress in a number of directions. In total numbers the membership is no larger but in revenue from dues there has been a marked increase. This has been due to the promotion to full membership of Junior Associates of earlier years, to the larger proportion that come in originally as members rather than juniors, and to the keener interest in the work of the Institute reflected in prompter payment of dues. The latter is important, since experience shows that of those who fall behind a large number must ultimately be dropped for non-payment. Conditions in the mining industries have been healthy though not booming. It has been much easier for members to find employment and meet their financial obligations but there has not been that flow of young men into mining, and hence in part into the Institute, that characterized certain earlier years. Rather the tendency toward centralization and larger industrial units has restricted the number, though not the scope, of opportunities for technical men in the mineral industries. The Institute is constantly receiving resignations from members crowded out of mining into other lines of activity. On the other hand, those who remain in mining and metallurgy have a greatly increased necessity for i,dormation and contacts with technology. A few years ago many superintendents and most foremen were non-technical men. Today the reverse rule obtains and shift bosses and men who occupy even lower positions in industrial units are called upon to know and to apply such an amount of technical knowledge that graduate engineers are the rule rather than the exception. The curious situation has obtained that with a much smaller enrolment of students in the schools of mining and metallurgy there has been at times an actual shortage of young engineers for beginners' jobs. The young men retort that the conditions of employment and pay offered in mining are not equal to those in other lines and, while it is undoubtedly true that the young engineer beginning in mining is able to start under conditions markedly better as to pay, living, and responsibility than did the men who are now leaders, they are required to know much more and have usually spent more in time and money to qualify for their jobs. The whole matter is one which warrants much more attention than has been given to it. For the Institute it is important since the young men

Jan 1, 1928

By R. M. Rose, D. P. Ferriss, J. Wulff

Single crystals of tantalum, tungsten, and the binary alloys thereof were grown by electron-beam zone melting and tested in tension between 77° and 373°K. The ductile-to-brittle transition temperature increased sharply with solute concentration, for both tungsten-rich and tantalum-rich alloys; apparently, the maximum transition temperature exists near the 50 pet concentration. Twinning was observed to be a consequence of cleavage-crack propagation. Cleavage occurred on(100)planes, and extensive tearing and "river markings" were present. The onset of brittleness could not be associated with any change in deformation mode, the yield point phenomenon, or the large rise in yield stress with decreasing temperature. Cleavage was observed to initiate at either the specimen surface, or subgrain boundaries. The observation was made that the grain boundaries of Polycrystalline materials are probably even better as nucleation sites for fracture, and are responsible for brittleness at high temperatures and lower solute concentrations than those of the corresponding single crystals. TUNGSTEN and tantalum, the two high-melting neighbors in the third transition period of the periodic table, present striking contrasts. Although both are bee, tungsten has far higher elastic moduli, and the highest ductile-to-brittle transition (DBT) of all the pure metals.' Tantalum, however, is ductile at liquid helium temperatures. Large strain aging and yield point phenomena are observed in tantalum,2,3 but only to a lesser extent in tungsten.4 Tungsten work-hardens at a high rate,5 whereas this effect is negligible in tantalum.' Since the mutual solid solubility of these two metals is complete,7 and no ordering has yet been reported in the binary system, the possibility of depressing tungsten's DBT by tantalum additions was considered. Also, tungsten additions to tantalum could produce hard, high-melting alloys which might be ductile at room temperature. Such alloys offered excellent bases for complex alloy development, particularly with regard to the effect of rhenium on Group VI metals.8"10 In addition, there was the opportunity to examine flow and fracture modes in the vicinity of the DBT, with particular emphasis on the correlation between brittleness and fundamental deformation mechanisms. Myers11 has reported that recrystallized alloys with up to 10 pet W in tantalum are ductile at room temperature, but alloys with 20 pct W could not be swaged at temperatures as high as 500°C. The whole spectrum of Ta-W alloys has been surveyed by Braun, Sedlatschek, and Kiefferl2 using sintered powder compacts. In addition to some physical properties they reported a maximum hardness at about 50 at. pet W, dropping to much lower hardness levels for the pure components. They were able to cold work tantalum-base alloys with as much as 30 pet W and to warm work tungsten-base alloys with up to 20 pet Ta by procedures similar to those used for pure tungsten. EXPERIMENTAL In order to study deformation modes and to keep the impurity concentrations at a minimum, electron-beam zone melting was used to prepare single crystals of high purity. The electron-beam zone melting apparatus is similar to that described by Calverley, Davis, and ever" and incorporates a grid control rectifier to maintain control of emission current. Pure metals were refined by making three zone passes at 4 mm per min up the rod. Alloys were prepared by melting together wire of one constituent on a previously melted rod of the major constituent. Subsequently, two zone passes were made in opposite directions to level the composition after the manner suggested by pfann.14 Purification of the starting materials was accomplished by evaporation of the several impurities, and the degree to which this was accomplished is indicated in the chemical analyses of Table I. The uniformity of composition along a zone levelled rod was within the i 1 pet accuracy of the fluorescent X-ray technique used to determine composition. Crystallographic orientation of the

Jan 1, 1962

By G. L. Oldright, F. W. Schroeder

Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may he considered that the improvements made were due largely to increases in the size of the furnace used. Doubtless some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By Norman Lamont

The evaluation of certain reser-voir properties, such as porosity and fluid saturation, from electrical well surveys has been widely accepted in petroleum engineering. Various investigators have established relationships between these properties and certain parameters which affect the response of the electrical log. Among these are the resistivities of the mud, its filtrate, and its filter cake. In 1949, Patnode1 established a relationship between the resistivities of the mud and filtrate. The well logging service companies have contributed relationships for the mud-mud cake resistivities2,3 These have been valuable since it was the practice to measure only resistivity of mud at the well site. During the mid-1940's the industry began drilling wells with oil-emulsion drilling fluids. These were conventional aqueous muds with a dispersed oil phase. Since 1950, oil-emulsion muds have been used on an increasing number of wells each year. However, the practice of measuring only the resistivity of the mud at the well site has continued, and the mud filtrate and mud cake resistivities have been determined by the above-mentioned relationships. Service companies are now equipped to measure all three resistivities at the well site. An investigation was conducted on the resistivities of oil-emulsion muds, mud filtrates, and mud cakes to determine if these values conformed to the relationships for aqueous muds. TYPES OF MUDS Fifty-one oil-emulsion mud samples were prepared in the laboratory following a standard manual' published by a leading mud company. The diesel oil in the samples varied from 5 to 50 per cent, the majority of the samples being in the 10 per cent region. The basic aqueous mud types which were converted to oil-emulsion muds were commercial clay and bentonite muds, low pH and high pH, caustic-quebracho treated muds, and lime treated muds. The emulsions were stabilized by dispersed solids, lignins, lignosulfo-nates, sodium carboxymethyl cellulose, or sulfonated petrolatum. It is worthy of note that after a quiescent period of two weeks at room temperature all samples, regardless of emulsifying agent, remained stable. The make-up water for the muds was from the laboratory tap. Resistivities were varied by the addition of table salt to the water. A range of mud resistivities from 0.44 to 3.9 ohm-m was obtained in this way. Twenty-three field muds were tested. These covered the same range of mud types as did laboratory muds. Oil provinces of the Gulf Coast, South Texas, West Texas, Oklahoma, Montana, and Canada were represented. MUD TEST PROCEDURE Each mud was tested for density, viscosity, pH, and filter loss by standard testing techniques. The resistivity measurements were obtained with a Schlumberger EMT meter. This meter required small volumes of sample, e.g., 2 mm. Filtrate was obtained from a Standard Baroid fil-ter press at the end of a 30-minute test. The filter cake from the same test was used for cake resistivity measurements. Mud, filtrate, and cake samples were heated to 100" F in a constant temperature water bath prior to measurement of resistivities. RESULTS The relation between mud resistivity (Rm) and mud filtrate resistivity (Rmf) is shown in Fig. 1. The solid line represents an average for the data. The equation of this line is Rmf =0.876 (Rm) 1.075 . . (1) Arbitrary limits, indicated by the dashed curves, have been set. The majority of the data falls within these limits, but some points do lie outside the limits. The approximate equation Rmt = 0.88 Rm , . . . . (2) will give satisfactory results within these limits. The data on mud cake resistivity Rmc is shown in Fig. 2. The solid line is an average for the data. The equation for the line is Rmc = 1.306 (Rm)0.88 The dashed lines are arbitrary limits on the data. Within these limits, Eq. 3 may be simplified to Rmc = 1.31 Rm . . . . (4) DISCUSSION The limiting curves in Figs. 1 and 2 represent maximum deviations of ±25 per cent. Thus the use of the average curves can introduce considerable error. There is no substitute for accurate measurements of mud, mud cake, and mud filtrate resistivities at the well site. The mud sample tested should be representative of the mud opposite the formation being logged. The average mud filtrate resistivity curve of Fig. 1 is reproduced in Fig. 3 with two curves which have been published for clay-base aqueous muds2,3. The latter curves were determined from average values of a large number of drilling fluids. The three curves have essentially the same slope and the differences between them are from 7 to 22 per cent. Comparison is made only to illustrate the possibility of error

Jan 1, 1958