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By D. J. Carney, R. L. Stephenson

M. O. Holowaty (Inland Steel Co., Chicago)—Messrs. Stephenson and Carney should be commended on the presentation of this very interesting paper. They have introduced new ideas on sampling of sinter and have thereby made a valuable contribution to the qualitative evaluation of sinter. The paper undoubtedly presents a novel approach to the problem of sinter sampling and this is to be regarded as its main value. Sampling of sinter from the pallets prior to the discharge is by no means new. This sampling procedure was used in the past by Philbrook; Hamilton and Ameen and others. The sampling basket used by the authors of this paper, however, permits one actually to obtain a good sample of the sintering mixture. The basket which can also be described as a removable pallet goes then through the entire process of sintering and thus offers a sample which is positively representative of the batch in question. The authors' method of evaluating physical properties of sinter seems to be not altogether clear, however. It is felt that the following points should be clarified: I—In the section describing the shatter test the authors state that a 5-lb sample was used in the actual tests. There is no mention of the initial size of sinter particles used and the method of obtaining this 5-lb sample from the sinter cake weighing about 100 to 120 lb. It is also surprising that the sample was dropped from a height of 3 ft instead of the conventionally used 6 ft. Unfortunately there is no mention of the number of drops to which the sample was subjected. 2—The tumbler test used by Messrs. Stephenson and Carney in testing the samples of iron ore sinter was performed in a standard ASTM coke tumbling drum. The drum measures 3 ft in diam and is equipped with two 3-in. lifter bars. The distance of fall in this testing drum is identical with the distance used in the previously described shatter test. The number of drops in the tumbling drum is approximately double the number of revolutions. The basic difference then between the drop-shatter test used by Stephenson and Carney and the tumbling test seems to be the number of drops to which the sample is subjected and the abrasive action to which the sinter particles are exposed in the rotating drum. In any case, the choice between the shatter test and the tumbler test seems to be only a matter of taste and convenience. 3—The authors of the paper state that "in order to insure getting results that were representative . . . the entire sample obtained by means of the sampling basket was used as the starting material for the tumbler test. If the sample broke in pieces, all the pieces were used regardless of their size." The initial size of the material is known to be a very important factor, influencing seriously the final size of the particles. This should be particularly important in case of hard sinters. It seems also very troublesome and inconvenient to charge 12-in. cubes of sinter weighing about 100 lb into the testing drum. This could be eliminated by taking a standard fraction of somewhat smaller particles.

Jan 1, 1954

By R. L. Perrine

The rise of drain holes drilled from a well to improve productivity is becoming increasingly common. This paper presents data on the productivity increase that may he expected. The data were obtained by use of an electrolytic tank mode1 of a reservoir. Parameter varied in the study to show their effect on productivity include the length, number, and arrangement of drain holes, formation thickness, and damage near the well-hore. The most important results are that very large increases in productivity can be expected only ill damaged wells. The distance, to which drain holes penetrate is of primary importance. Additional length is more effective than increasing the number of drain holes. INTRODUCTI O N In planning the use of drain holes to improve productivity, it is important to he able to estimate the increase in production that can be obtained for a given expenditure. It is also important to know what combination of parameters related to the use of drain holes can he chosen to give the largest production increase. This paper presents data on the productivity increase that may be expected, as obtained by use of an electrolytic tank model of the reservoir. In this study, :! number of parameters were varied to show their effect on productivity. These included the length, number, and arrangement of the drain holes. The presence or absence of a damaged zone of low permeability around the well-bore and the thickness of the producing formation were also considered. The productivity increase that might be expected due to drain holes is presented as the ratio: productivity of a well after drain holes are drilled and completed to the productivity of the same well in an undamaged state without drain holes. The symbol Q/Qo is used to represent this productivity ratio. Thus, the comparison in each case is with a formation of uniform and undamaged permeability. Additional notation includes I;, for the radial extent of the damaged zone, which is assumed to be cylindrical and concentric with the wellbore. Thus the effect of an actual block is approximated by a discontinuous varia-tion in permeability. For purposes of this paper, the permeability within the damaged zone, denoted by kb, is assumed to be 1/25 that in the remainder of the formation, or k. A necessary assumption in making comparative tests using an electrolytic model is that reservoir flow behavior can be approximated by single-phase, steady-\tate flow according to Darey's Law. In addition, the formation must be assumed to be homogeneous with equal horizontal and vertical permeability. The well-hore diameter in this study is assumed to be 11¾ in. and the drainage radius of the well equal to 500 ft. The drain holes are considered to have been completed without damage. THE ELECTROLYTlC MODEL 'She producing formation was modeled by forming a sheet of phosphor-bronze into a cylinder 40% in, in diameter to represent the boundary of the drainage area. This was enclosed in a waterproofed plywood box, and the box filled to an appropriate level with a dilute sodium chloride solution. Solution concentration was adjusted to give a total resistance from

Jan 1, 1956

By E. W. Davis

ABOUT a year ago several carloads of magnetite ore from New York state were sent to the Mines Experiment Station at the University of Minnesota for grinding and concentration tests. The flowsheet used minus ½ -in. ore as feed to a 3 by 3-ft. overflow cylindrical ball mill operating at 37 r.p.m. in closed ciruit with a vibrating screen. The screen oversize was returned to the mill for further grinding and the screen undersize was concentrated magnetically. Several different screens, between 8 and 35 mesh, were used and it was desired to determine the capacity of the ball mill operating in closed circuit with each screen. From earlier tests using this flowsheet and equipment, it was known that when comparatively coarse screens were used it was necessary to increase the grinding dilution in the mill beyond that ordinarily used when grinding in closed circuit with classifiers. It was also known that after the circulating load had passed about zoo per cent of the original feed it tended to increase rapidly in amount, and grinding conditions appeared to become unstable. These conditions may be peculiar to this small ball mill but they were very marked in this series of tests. If conditions were not right, the mill would suddenly "go dead," and it was found desirable to operate so that the occasional rattle of balls against the lining of the mill could be detected. As the test proceeded, it was found that the cure for almost any difficulty that developed in the grinding circuit was to add more water to the mill feed. To begin with, the mill was operated at pulp densities of from 60 to 65 per cent solids but gradually it was found that with higher dilutions grinding conditions were more stable and the ball-mill power per ton of screen undersize decreased. Actually, the most satisfactory operating conditions were secured with a pulp density of about 40 per cent solids. The ore was hard but it was possible to grind ½ in. feed so that all would pass an 8-mesh screen with a power expenditure in the ball mill of less than 2 1/2 kw-hr. per ton and with a circulating load of less than 50 per cent. With so high a dilution in the feed to the mill, it was feared that steel wear would be very high. However, the mill was not noisy, and in order to find exactly the grinding conditions, a special test was made at the end of the main series of tests. After the mill had been in operation for several hours and grinding conditions had become balanced, the power to all the equipment was shut off at a given signal. The ball mill made about one revolution after the power was turned off, and just as soon as the mill stopped, a spout was placed at the discharge trunnion and all of the pulp that flowed out of the mill was caught in tanks. The manhole in the mill was then opened and the whole mill charge dumped into other tanks. Both of these products were weighed, dried, and weighed again. The balls were screened out and weighed, and the dry solids were sized. In this way it was possible to determine the nature of the pulp and the ore within the mill under actual grinding conditions.

Jan 1, 1945

By E. V. Potter, H. C. Lukens

In general, during all electrowinning processes, large volumes of gas are liberated at the cathodes of the electrolytic cells. Most of this gas escapes from the electrolyte, but much of it may be absorbed or entrapped in the metal being plated. In many cases, this is not objectionable, but since it is liberated when the metal is heated, it is a potential source of trouble where the metal has to be heated or melted. Since this gas is principally hydrogen, in escaping from a melt it may cause minor explosions and scattering of the molten metal, while in other cases it has been reported to make the metal rise in the molds when cast. Therefore such metals would be more generally useful if all or most of the gas were removed before they were used. In an earlier investigation1 the gas content of electrolytic manganese was determined and methods found for removing most of it without detrimentally affecting the properties of the metal. In this paper results of a similar investigation on electrolytic chromium are presented. Apparatus The apparatus used in these investigations is essentially the same as that used in the manganese work1 and is shown in Fig I. It consists of sample chamber A, vacuum or pressure gauge B, volume gauge C, vacuum pump D, hydrogen container E, auxiliary gas chamber F, and furnace G. The pressure in the system is measured by the mercury manometer B. Changes in gas volume are measured by volume gauge C which is essentially another mercury manometer in which the volume of the mercury column is large and can be varied by raising or lowering the well H. In this apparatus the tube had a volume of 100 cc for a 60 cm length. Larger changes in gas volume can be accommodated by auxiliary containers attached to the system at F. The sample is heated by furnace G, the temperature of which is regulated within 5°C by an automatic controller. The only change in the apparatus was the sample container which was made smaller because of the chromium metal's containing more gas than the manganese metal and a smaller sample was used. Method The measurements were made as follows: The complete system when cold was evacuated and refilled with H2 until all the air was removed. It was then refilled with H2 to the desired pressure, as indicated on manometer B. The temperature was then raised to the desired value and as gas was either liberated from or absorbed by the sample, the height of well H was changed to alter the volume of the system so as to keep the pressure constant. In determining rates of liberation of gas, volume readings were taken at convenient tirne intervals depending on the rapidity with which the gas was liberated. In determining total volumes, the specimen was held at constant temperature and pressure until

Jan 1, 1949

By T. M. Geffen, F. F. Craig

Laboratory model flow tests have been made to simulate field conditions of partial pressure rnaintenance by dispersed gas drive on rocks having sandstone-type porosity. In this production method there is a significant saturation gradient between injection and production wells. A means of calculating relative permeability characteristics. from partial pressure rnaintenance flow tests has been developed. This method takes into account the saturation gradient. These kg/ko characteristics are identical to those for solution and external gas drives. This method of calculation, which has been proved experimentally, has limitations; however, it is valid for cases likely to be encountered in field prediction of partial pressure maintenance performance. The production performance of a hypothetical reservoir has been calculated for several different injection programs, involving both full and partial pressure maintenance. Results indicate that the maxinum oil recovery by these gas injection programs is obtained by full pressure mainte- nance at the original bubble point. Correspondingly lower oil recoveries are obtained by postponing gas injection and/or by reducing the degree of pressure maintenance. INTRODUCTION Dispersed gas injection is an important method of recovering additional oil beyond that obtainable by pressure depletion alone. The injected gas acts as a source of additional energy to displace oil from the rock formation as well as to retard oil shrinkage in the reservoir. In practice, gas injection is usually started after the reservoir pressure has fallen below the original bubble point. For various reasons sufficient gas is generally not injected to completely maintain the pressure in the reservoir. So, injecting gas at a declining reservoir pressure results in a combined external and solution gas drive. These two oil displacement processes are characterized by different gross saturation distributions in the reservoir. The external gas drive, like all frontal drive processes, inherently results in an oil saturation which increases from the injection wells to the producing wells. In the solution gas drive there is essentially a uniform oil saturation throughout the reservoir, provided the field is being depleted uniformly. It has been customary to predict oil recovery performance under partial pressure maintenance by a method which does not take into account the saturation gradient1. It has been shown2 that neglecting the presence of a saturation gradient can cause considerable error. It is the purpose of this paper to develop and present an improved method of predicting partial pressure maintenance performance of reservoirs including, where necessary. the effects of well arrangement. From the laboratory standpoint, this problem involves the measurement of the basic flow behavior (relative permeability characteristics) under partial pressure maintenance conditions. Recent research- as shown that for rocks having sandstone-type porosity, the relative permeability characteristics are identical for solution gas drive and complete pressure maintenance external gas drives. Since both solution and external drives have the same basic flow behavior, it was speculated that partial pressure maintenance, being a combination of the two, would also

Jan 1, 1957

By Leon W. Dupuy

Many articles have been written describing peculiar and particular types of drilling. Little correlation has been made between the character of ground to be drilled and sampled and the type of drilling most desirable for procuring satisfactory samples. In development work, the engineer and metallurgist often desire the recovery of a sample in as physically natural a condition as possible, particularly when the sample must provide material for process tests as well as chemical and physical analyses. This paper describes the results of experimentation in drilling unconsoli-dated materials carrying titanium minerals in the Magnet Cove area, Ark. The drilling was most difficult in that the ground varied much in consistency, and it was vitally necessary to recover a physically natural sample. Ordinary churn drilling recovered a sample that could be analyzed chemically, but the sample was not in its natural state. The problem was successfully solved by utilizing and further developing methods used in the oil and clay industries. In the Magnet Cove area of Hot Spring County, Ark., the titanium minerals include rutile, brookite, ana- tase, leucoxene, and others. They occur in soil, clays, and altered rocks adjacent to novaculite in contact meta-morphic deposits. In preparing titanium minerals for use in the industry, it is necessary to know the physical characteristics of the ore. Also it is believed that the particular physical character of the titanium mineral used is just as important as the chemical character. Brookite and rutile, for example, are the same chemically but differ in crystalline structure. In preparing titanium minerals for market, beneficia-tion is necessary. For the millman to make a satisfactory laboratory test, the sample he uses must be as nearly of the same physical character as he would have from mine run ore. In the past at Magnet Cove, churn drilling had been done on the properties, and assays of total titania had been obtained. It was found, however, that the total titania was indicative merely of the presence of titania but did not tell the metallurgist the character of the material or even give him the slightest indication of what he might expect to recover. When the Bureau of Mines entered the field in the winter of 1947 and 1948, it was therefore necessary that a natural sample be recovered. LIMITATlONS IN ORDINARY PRACTICE The churn drilling that had previously been done had shown that the occurrence of titanium-bearing mineral was somewhat widespread. The samples were, however, completely beaten up and in a finely powdered state. The minerals had been slimed, and such ore dressing tests as were made from the churn-drill samples were found to be worthless. Test pitting and outcrop sampling had provided satisfactory samples of surface ore but did not represent the entire ore body. A quantity of rutile had been produced from the surface residual concentrations in the westerly part of the cove, but production ran into metallurgical trouble as depth was attained. Representative deep samples upon which ore dressing research might depend were needed.

Jan 1, 1950

On Friday, Jan. 17, the thirteenth impression of the Perkin Medal was presented to Dr. Frederick Gardner Cottrell, at a meeting of the New York Section of the Society of Chemical Industry, held at the Chemists' Club, in New York City. The presentation address was delivered by Dr. C. F. Chandler, Senior Past President of the Society. The medal was awarded for the recipient's work in the electrical precipitation of suspended particles in gases. Dr. Cottrell's career, which is so familiar to members of the Institute, was sketched by Dr. Chandler. He told at some length of the experiments leading up to the final solution of electrostatic precipitation, which has done so much to solve the problem of smelter smoke. The principal interest in Dr. Cottrell's speech of acceptance lay in his remarks about helium, the production of which has been so reduced in cost that it has practically supplanted the use of hydrogen for inflation purposes in war balloons. He said: "Two or three weeks after war was declared in 1917, Dr. R. B. Moore attended a meeting of the American Chemical Society in Kansas City. At the general session, Mr. Seibel of the University of Kansas, gave a paper on krypton and xenon in some of the natural gases of Kansas. At the end of his paper, he expressed regret that at such a time, when everyone was thinking of war problems, his paper was of a purely scientific nature and had no practical bearing on the war. Dr. Moore immediately got up and said that he did not agree with Mr. Seibel. The presence of helium in those wells could, and should, have a very practical bearing on the war, as this gas could be extracted in quantity from the natural gas and used for balloons and Zeppelins. He pointed out some of the natural advantages of helium over hydrogen, and quoted Sir William Ramsey's letter advocating the use of helium for airships. The general attitude of those present was one of scepticism, so he finally asked Dr. Cady for his opinion who stated that he believed-the thing could be done, but was very doubtful whether it could ever be made practical on account of the cost." Dr. Moore in a letter wrote: "That same day I talked to Dr. Parsons, who was present at the meeting and was returning to Washington almost immediately. I told him what had happened, and that I believed the matter should be taken up by the Bureau at once and presented to the War Department. He promised to do this as soon as he got to Washington; and I know that he did take the matter up with other Bureau officials." Dr. Cottrell continued in part: News of Ramsey's suggestions also reached this .country through other channels and eventually came to our attention. Professors Satterly and Patterson of the University of Toronto commenced experimentation of the subject Jan. 1, 1916. Col. G. A. Burrell, who headed the research department of the Gas Warfare Service at the American University, tells me that Cady's experiments had also suggested similar possibilities to him.

Jan 3, 1919

By H. W. Schreiber, G. R. Castle

INTRODUCTION The small mining company faces the dilemma of how to finance the development of its properties. Many of these small companies don't have the financial resources to pay for development costs from their own funds and frequently they do not even have the financial strength to permit them to borrow the money needed for development. A project financing can sometimes be arranged in such a way as to solve this dilemma. A typical project financing will not provide the answer to the "little guy's" problem because most project financings require that the sponsors of the project provide completion undertakings and sometimes other undertakings. These undertakings are often a prerequisite to obtaining financing, and if the sponsor is not financially strong enough to provide the comfort that the lenders require, then the sponsor will have difficulty obtaining financing no matter what undertakings are offered. However, with a little imagination and the right set of circumstances, a project financing can sometimes provide the solution to the small mining company's problem. The Strathcona Sound Project is unusually suited as a case study because the sponsor, Mineral Resources International Limited,(MRI) was able to arrange a project financing for the development of the project even though the lenders probably would not have been willing to loan a similar amount of money direct to MRI based on the full faith and credit of MRI's balance sheet. The way that MRI was able to arrange this financing makes for an interesting story. THE CASE In early 1974, MRI owned 100% of a small to medium sized zinc - lead - silver massive sulfide deposit (Property) located 475 miles north of the Arctic Circle less than two miles from deep water on Strathcona Sound which is a shipping zone in the Canadian Arctic. The Property was situated at a desolate location except for an Eskimo settlement (Arctic Bay) which was about 20 miles west of the Property. In other words, the project required that complete infrastructure would need to be installed in this far northern location. Independent reserve and feasibility studies (Study) had been completed on the Property which confirmed a total of 6,825,000 tons of reserves classified and summarized as shown below on Table I: [ ] These reserves were based on 229 diamond drill holes totalling 58,909 feet and were established on a cut-off grade of 7% zinc and a minimum mining height of seven feet. The overall mining recovery rate was estimated at not less than 85% and not more than 90%. The feasibility study indicated that it would cost an estimated $45,044,000 to develop an underground mine, a concentrator, a town-site, a power plant, storage facility, loading dock, a tailings disposal area and an airstrip on the Property with an annual capacity of 525,000 tons of ore per year. The construction period was expected to take two years and projected capital costs were broken down as set forth below on Table II:

Jan 1, 1985

By N. A. McLeod

BALL wear in wet grinding mills has been the subject of considerable discussion in the last few years. Its importance to millmen is obvious in view of the fact that ball wear may cost from 2 to 4¢ per ton of ore for the more fortunate operators and many times this cost when a refractory ore must be reduced to a fine state of comminution. Consequently, any factual information that may be obtained on the nature of ball wear may be applied to introduce changes leading to lower ball consumption and therefore also to lower costs. In 1940, while employed with the Britannia Mining and Smelting Co., Britannia Beach, B.C., I had occasion to investigate factors that might affect ball consumption. We performed a number of tests, which will be briefly described. These tests originated as a result of changing from a 3 1/2-in. ball to a 2-in. ball of different type of manufacture and were motivated by the desire to obtain the best possible type of 2-in. ball. Our testing was helped by the fact that the company operated a one-ton electric furnace, so that it was quite easy to make steel changes. It should be emphasized at the beginning that these tests produced results that in their general application to milling should be interpreted qualitatively rather than quantitatively. As will be evident in the description of the method used in testing, the balls were subjected to an intermittent batch grinding treatment, which may have produced results in rate of steel wear that may not have an accurate correlation with steel consumption in full-scale mill operation. However, with this fixed and rather arbitrary method of testing, results were obtained that may be of some interest and may stimulate further work on this subject. Unfortunately, owing to pressure of war work, our tests, which we had intended to repeat on a continuous grinding operation, had to be discontinued. METHOD OF TESTING Tests were made in a small cylindrical mill, rotating at 72 r.p.m. Internal dimensions of the mill were 10 1/2-in. dia. by 9 in. The cylindrical inner surface was lined with rubber and the two ends were formed by steel plates. A charge of 2000 grams dry feed and 2000 c.c. of water was added every 2 hr., the mill being drained of pulp before the addition of each charge, by tilting and sluicing with water. The mill was run on afternoon and night shift for daily periods of about 15 hr. total grinding time, and each test lasted from 60 to 150 hr. This batch method of testing involved a higher average pulp temperature in the mill, which undoubtedly affected comparative results in some of the tests. HARDNESS OF BALL VS. WEAR In these tests, special "eutectoid" steel balls were used, some of which were chill-cast and quenched, some sand-cast and quenched, and some sand-cast and annealed to a dead soft condition. The results obtained, using as mill feed an ore containing considerable free quartz and

Jan 1, 1945

Book Review MEXICO UNDER CARRANZA. By Thomas E. Gibbon, Los Angeles, California. Doubleday Page and Co., New York, 1919, 270 pp., 711/2 X 5 in. $1.50. A vivid, accurate, convincing summing up of the Mexican problem as it confronts us today, devoid of sensationalism or bigotry. A compelling narrative of the relations of the Carranza Government and our own from Oct. 15, 1915, when it was recognized as the de facto government of Mexico, to the present time. It shows the absolute lack of promises kept or lived up to either with its own people or the United States. This is not done by invective or the indiscriminate use of adjectives but by quoting from the instruments on record on both sides of the case. The driving of the Red Cross out of Mexico, which condemned many of the native peons to starvation, and the interference with the Pershing Punitive Expedition are merely mentioned, but the crying crime against the Mexican people, who have sunk to unbelievable depths of misery obscured, but not exceeded, by the sufferings in Europe, is held up in a clarified light. Two illuminating chapters are devoted to the character and treatment of foreign investments in Mexico, Which are a complete refutation of the common statement that Mexico has been exploited remorselessly by the foreign investor. The conces-sions given by Diaz in the days of his regime most condemned are nowise as large nor the land grants as extensive nor the title so absolute as those granted to the railroad builders of our own transcontinental systems and other of our nationally subsidized works. This is not left to further argument but is settled by quoting the various official instruments and not at all in a tedious way. The author's concluding chapter is a picture of the Mexican people from Aztec times, when they had reached a civilization which gave them a better calendar than the Roman Empire ever had until modern times, to their present unhappy and pitiful condition. He concludes, basing his findings on parallel historical downfalls of a once proud people, that present conditions are the natural result of the absolute domination of an alien minority. He grants, of course, that there are many estimable Latin-Mexicans of integrity, ability and charm, but maintains that the dominant minority has continually degraded in fulfillment of the sociological law that the sure revenge of the servile class is found in the corruption of the master class. The book is replete with historical and literary allusions which, added to the pleasing style of the accomplished author, make fascinating the account of one of the world's great tragedies. Mr. Gibbon is a lawyer who has spent much time in Mexico and shows an understanding of the native people not shared by all writers on the Mexican problem. A distinct addition to all literature on the subject. P. E. B.

Jan 8, 1919

By Irving M. Symonds

COMPAÑIA Minera de Peñoles, S. A., at its Avalos unit operates a lead-zinc concentrator having a capacity of about 600 short tons per day. Lead and zinc concentrates are made by flotation methods. The concentrator is at Terminal de Providencia, in the state of Zacatecas, some 70 miles south of Saltillo, Coahuila. The Peñoles-Avalos railroad operating between the plant and Estacion Avalos, a distance of 6 miles, makes connections with the Coahuila and Zacatecas railroad, which in turn connects with the Nacionales de Mexico at Saltillo. The three railroads are each of different gauge, requiring the transfer of all supplies and concentrates at each junction. The mill operates on company ores, on ores from leased mines, and as a custom mill. The custom business is very small. The lead concentrates are shipped to company smelters at Monterrey, Nuevo Leon and Torreón, Coahuila. Since the European war, all zinc concentrates have been shipped to the United States. ORE TREATED The major source of ore supply is a group of four mines owned or leased by the company, some three miles south of the mill in the mountains, from which the ore is delivered to the mill by a gravity aerial tramway. Custom ore comes in by rail. The ores occur in chimneys. The upper levels contained oxidized siliceous ores of lead, copper, silver and gold, with varying amounts of zinc. These ores were direct-smelted. In the first two years of operation a good deal of the tailing loss was caused by the oxidized content of the ores milled. Much of the ore in the transition zones between the oxides and sulphides was not mined. The ores now being milled are clean, hard sulphides in a limestone gangue, with only occasional contamination of oxides. The lead and zinc sulphides are of coarse structure and are freed even above flotation sizes. The galena is usually "steel" colored, easily and quickly floated. When there is any oxidation, it is usually on the coarsest crystals. The zinc minerals vary a good deal in color; a resin sphalerite predominates but there are varying shades of brown marmatites. On an average, the zinc sulphides contain 64 per cent zinc. The only common copper mineral is chalcopyrite. Little is known of the occurrence of gold and silver. Waste minerals are mainly limestone and pyrite, both of which are barren. The ores vary from 20 to 35 per cent pyrite. Mill heads for the year 1940 and the first eight months of 1941 are given in Table I. [ ]

Jan 1, 1942

By B. Otto Pixler

Mud-analysis logging provides identification of the fluid content of formations drilled by the rotary method and permits the accurate correlation of oil and gas shows with the depth. Mud-analysis logging eliminates much coring and has an especially important application in drilling programs when coring is particularly hazardous or the electric log is difficult to interpret correctly. An appreciation of the factors that influence the results of the mud analysis, such as rate of penetration, mud properties, coring, and characteristics of reservoir rock as related to the "flushing effect," permits a better understanding of the mud-analysis log. Recent studies of the relationship of the two gas readings obtained by the gas-detector instrument have proved significant. The difference between the total combustible gas reading and the higher molecular weight gas reading indicates the presence of methane. With few exceptions, zones not showing an increase of methane probably can be condemned. Mud-analysis logging furnishes valuable information, which generally cannot be obtained by other methods. Introduction Logging by mud analysis of wells drilled with rotary tools is a direct method of locating oil-and-gas-bearing formations. The use of this method of logging has increased each year until now more than 20 field units are being used in 13 oil-producing states and in some foreign countries. Several papers1-9 have been published in the petroleum trade journals describing in detail the equipment and the application of mud-analysis logging. This paper will discuss briefly the method and review its application and some of the late developments. Method As the drilling of a well progresses, the drilling bit dislodges and disintegrates a cylindrical section of the formation. If oil or gas is contained in the pore spaces of the cylinder, some of the contents of these porous spaces will be entrained in the drilling fluid. In wall logging by mud analysis, the mud is continuously examined for oil and gas on its return to the surface, and the results are correlated back to the actual depth of the hole at the time the mud was clearing the eyes of the bit. The test for gas is made by diverting a portion of the circulating mud from the flow line to a separator or gas trap, where the mud is thoroughly mixed with air and a portion of the gas entrained in the mud is removed. A stream of air is drawn coun-tercurrent to the flow of mud in the gas trap, thus materially assisting the separation of the gas from the mud. The air-gas mixture is then drawn into a "hot-wire" gas-detector instrument where the percentage of the combustible gas is determined. The presence of oil in the formation drilled is detected by a physical examination of the drilling mud under ultraviolet light. A sample of the mud is treated to reduce the surface tension and gel strength, after which it is placed in a viewing box. This box is so constructed that all external light is excluded and so that the sample may be subjected selectively to either ultra-

Jan 1, 1946

By L. S. Darken

IT has been known for sometime that in an ionic lattice, such as that of Ag2S or FeO, the migration velocity of the anion may differ markedly from that of the cation, the cation being usually the more mobile. It has been shown, for example, that during the formation of silver sulfide tarnish on silver, the tarnish layer grows principally by the motion of silver cations through an essentially stationary silver sulfide lattice. Similarly, the growth of wüstite scale on iron proceeds principally by motion of iron ions, this being accomplished with the aid of vacant iron lattice sites into which the iron may readily move. Likewise investigations of solid solutions of various salts show that there is usually considerable difference in the mobility of the different cations. Measurements of the mobility of various metals in liquid mercury indicate that the various metals move at different rates relative to the mercury. In view of the foregoing it seems reasonable to suppose that when any single phase solution is acted upon by a field, the atoms of different elements respond in different ways; in particular, that the force arising from a particular concentration gradient in a binary alloy would cause the atoms of one component to move with a drift velocity which is in general different from that of the atoms of the other component. The recent experiments of Smigelskas and Kirkendall1 support this viewpoint, since they found, in essence, that during diffusion in alpha brass zinc moves appreciably faster than copper. This was evidenced by the relative motion of inert molybdenum wires. Such a phenomenon is incompatible with the place interchange theory of diffusion proposed by several authors, most recently by Birchenall and Mehl.2 This theory has, however, also been questioned3.4 for other reasons, principally because of the high energy required for direct atomic interchange. The rather large difference between self-diffusivity, as determined by tracer technique, and mutual diffusivity, as determined with the aid of another element at low concentration, might also be interpreted as favoring the above idea; however, Birchenall and Mehl have pointed out that there is another possible interpretation of this difference; this is discussed later. A rather interesting and pertinent observation was made by Schwartz,5 who noticed that tungsten lamp filaments, heated by direct current, commonly fail by necking down in the region at the temperature gradient near the positive end, with a concomitant swelling at the corresponding position near the negative end. He introduces considerable auxiliary evidence to show that this phenomenon is occasioned by the motion of tungsten cations which migrate in the high temperature region from the positive end toward the negative end. The mechanism of this migration is not clearly understood. If, however, this experimental evidence be accepted, we are forced to conclude that alloying elements in the tungsten may migrate by the same or a similar mechanism, whatever it may be, at a greater or lesser rate, thus giving rise

Jan 1, 1948

By E. W. Gieseke

Coal preparation men today are confronted with problems brought about by changing feed characteristics. The size consist of the coal going to preparation plants has been getting finer and finer. In order to handle these sizes economically, operators are turning to Notation. American Cyanamid Co., in an effort to enable coal flotation plants to produce efficiently at maximum quality and capacity, wnducted experiments which demonstrated that synthetic flocculants are useful for increasing the capacity of coal and refuse filters and for clarifying coal and refuse slurries, and a wetting agent is beneficial in reducing the moisture content of filter cake. Coal preparation men are today confronted with problems brought about by the changing coal markets, changing feed characteristics, and ever increasing labor and supply costs. Small mines are finding it difficult to exist, and the larger mines are rapidly converting to fully mechanized and automated operation to survive. The size consist of the coal going to preparation plants has been getting finer and finer, and the problems facing the preparation plant personnel are growing larger and larger. Facilities can be provided to handle this plant feed in new plants, but the problems are generally not so simple in existing plants where equipment must often be used to perform tasks far beyond its rated capacity. In addition, various state conservation departments, water commission boards, and other Federal and state agencies are endeavoring to eliminate stream pollution. These problems, and many more not mentioned, have been forced on the preparation engineer for reasons of economy and the public interest. In the matter of stream pollution, a determined effort is being made by the coal industry to cooperate with the Federal, state, and local agencies. Economists predict it will be necessary to produce 600 million to 700 million tons of coal by 1970. Bear in mind that the greater proportion of this increased tonnage will have to be processed in existing plants, and you get an idea as to the problems facing the preparation engineer. Methods presently in use for preparing fine coal do an acceptable job down to 48 to 100 mesh. A very large percentage of the - 48 or -100 mesh fraction of the tonnage of coal produced now is being sent to settling ponds or sold at a low figure for reason of quality. This has caused operators to turn to flotation. Building settling ponds and impounding fines is a costly and, at times, impossible method of handling these sizes. Recently a number of flotation plants have been installed to beneficiate material finer than 48 mesh, and these plants have demonstrated that flotation can be successful and increase the plant realization. Coal is a low cost commodity and processes for its beneficiation must be low in cost. Flotation is such a process. It is easily adapted to handling fine coal, requires very little supervision when properly installed, and is economical to operate. Flotation separates coal from refuse and fine clay. For reagents, several combinations are used. Some

Jan 1, 1962

By Walter Miller

During the second year of America's active ia in the war the inain objectives of the petroleum-refining industry were again to provide the four most important product needs for war: 100-octane aviation gasoline, toluene for T.N.T. production, high-quality lubricating oils for the needs of aviation and the armed forces, and petroleum chemicals for synthetic rubber manufacture. All other materials supplied by the petroleum industry took an inferior priority ranking, depending on the degree of essentiality. To accomplish the above objectives the industry initiated and carried on a building program during 1942 and 1943 which will put those two years down in petroleum-refining history as the greatest construction years of all time, a record which will hold for decades to come. Led by an estimated total of $ 900,000,000 going into 100-octane and contributory plant investment, it is easy to picture a total of two billion dollars and more going into additional refining-industry plant investment during 1942 and 1943, most of it privately paid for, but a substantial part covered by R.F.C. Government money financing. Some of the new plants will not be completed until 1944, but a large proportion of the program was finished and put into operation, adding substantially to the output of the four principal products during 1943- lne most important and spectacuiar activity of the refining industry is of course the 100-octane gasoline program necessitated by the fact that nearly 90 per cent of the requirements of the United Nations is being furnished by the United States. Recent P.A.W. figures mention $goo,-ooo,ooo as the cost of the construction looking to increasing the output of this all-important munition. Thirty-four major plants have been completed, 11 of them during the last quarter of the year. Thirty-eight more will be completed and put into operation during the first four or five months of 1944. Twenty-two additional plants authorized during 1943 are already well under way and scheduled for production late in 1944. Early in 1942 the aim was to treble the capacity for making 100-octane gasoline then existing; later, plans were to make it five times as great, to match the increased warplane building program; and toward the end of the year the sights had been set at a still higher but unannounced target, representing six to seven times the early 1942 figure. Comparative figures issued at the end of 1943 enable one to guess a program by the middle of 1944 of eight to nine times the former production, with an ultimate production in 1944 of eleven or more times the early 1942 output. It became obvious in 1943 that building of new units would be materially slowed because of delay in getting construction materials and that they would reach operation at considerably later dates than

Jan 1, 1944

By F. S. McNicholas

IN the writer's opinion the term "block caving" has been rather loosely applied to a general principle of mining, and therefore detailed classification of caving methods and draw, together with the characteristics required for each method, will fill a desired need. Selection of caving method should be made only after a study of all governing conditions, such as size of orebody, lateral and vertical extent of ore-body, attitude, treatment demands, possible resultant pressures, and cost and profit comparison of methods. Successful caving methods can be classified under six main headings: (1) Full-protected block caving, (2) semi-protected block caving, (3) unprotected block caving, (4) full-panel caving, (5) half block-panel caving, (6) block-panel caving. The methods and factors governing choice of method are shown in Table I. Full-protected Block Caving Full-protected block caving consists of develop-ing, undercutting, and drawing to exhaustion a block of ore of predetermined size without coordination of mining or draw with adjoining areas. The block is mined and drawn as an independent unit (Fig. 1). Full-protected block caving demands that mining and draw cannot be done against unconsolidated capping in adjoining caved and exhausted blocks. Two sides of block are against solid. Caved capping must be allowed to consolidate for a suitable period, usually not less than one year. Pillars may be left in lieu of consolidated capping (Fig. 2). The advantages of full-protected block, caving are: (1) A definite area is developed, mined, and drawn as a unit so that all work may be completed, the block drawn to exhaustion in the minimum length of time, and the area then abandoned. (? Production is not restrained because of coordina1, on with adjoining areas. (3) Records of draw and caved reserve of ore are good. Upon abanwnment of the block, accountability for production is closed out with no carry-over to adjoining areas. The disadvantages of full-protected block caving are: (1) Side dilution from the draw adjacent to the consolidated capping may be serious, particularly where capping does not consolidate sufficiently in the allotted time. (2) Boundary weakening is needed on the solid sides of the block except in weak rocks. (3) Maximum cantilever and compression effect to aid caving and fragmentation may not be secured since the draw should proceed from the solid to the consolidated capping to minimize side dilution. However, undercutting from the loose (caved capping) to the solid with a slight initial draw in the same manner will develop cantilever effect. (4) A period of at least one year, more is usually desirable, must be allowed for caved capping ,to consolidate before mining the adjoining block, depending upon the characteristics of the capping. (5) Capping may not consolidate sufficiently in the time that can be allowed. Full-protected block caving is indicated where: (1) The ground is weak and heavy and workings

Jan 1, 1952

By J. H. East, E. R. Maize

INTRODUCTION PROBABLY no other type of equipment is now being introduced into American underground mines about which less is known and about which there is more misinformation than the Diesel mine locomotive. Any equipment installed underground brings with it hazards that were not present before, and it is the purpose of this paper to show that Diesel-powered equipment can be used with safety in underground mining if certain feasible but very important precautions are taken. Diesel locomotives are not new in underground mining, since they have been used successfully in European mines for at least 19 years, but their use in American mines is exceedingly limited; so far as is known, only five Diesel mine locomotives are operating at present in United States mines. Diesel locomotives were permitted underground in Great Britain for the first time in 1939, and the rapid increase in the number in use in those mines, testifies to their safety and efficiency. More than 1300 Diesel mine locomotives are in use in European mines, according to the latest statistics; these locomotives are employed in coal, iron, lead-zinc and mercury mines. So far as is known, the Diesel mine locomotive has not yet caused an explosion or mine fire, and has not been the cause of any fatalities from toxic gases in the mines in which it has been used. ADJUSTMENTS OF DIESEL ENGINES FOR UNDERGROUND USE A Diesel engine adjusted to work in an underground mine should not be confused with the same engine adjusted for use outside the mine. Adjustment of the air:fuel ratio for engines that are used on the surface is intended to give the maximum fuel economy, and little or no attention is given to the chemical composition of the exhaust gas produced. Many such engines are poorly adjusted and perhaps poorly maintained, with the result that the engine emits smoke in quantity as well as the disagreeable odor familiarly associated with Diesel engines. The air:fuel ratio in the case of Diesel engines operated on the surface is generally around 15:1; that is, 15 lb of air to burn 1 lb of fuel. Diesel engines adjusted to work underground usually have an air:fuel ratio of about 20:I, as at this ratio the minimum quantity of toxic gases is given off in the engine exhaust. Diesel engines operated underground must be maintained in better condition than similar engines used on the surface; otherwise, the smoke emitted in the exhaust becomes objectionable if not unendurable. Practically normal air is required for safe operation of Diesel engines underground or in confined places; failure to provide virtually normal air for Diesel engine operation is almost certain to result in trouble. Bureau of Mines engineers have determined that the air in which a Diesel operates in confined places should contain less than 1/2 pct of carbon dioxide and at least 20 pct of

Jan 1, 1947

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By C. L. Prokop

A laboratory investigation has been made of the effects of mud hydraulics upon the formation and erosion of mud filter cakes. The tests were conducted to simulate drilling conditions as nearly as possible. The formation of mud filter cake in a drilling well does not proceed at a uniform and unbroken rate. Instead, the rate of cake accumulation depends upon whether or not the mud is being circulated. If the mud column is quiescent, filter cake formation is a smooth function of the filtration characteristics of the system. If the mud is being circulated filter cake formation depends not only upon the filtration characteristics of the mud but also upon the erosive action of the flowing mud column Filter cakes formed during continuous mud circulation were observed to reach an equilibrium thickness after several hours' circulation. Mud circulation was maintained at a constant volumetric rate throughout each experiment. The fluid velocity at equilibrium cake thickness was dependent upon the thickness of the filter cake. Muds having exceptionally high water loss deposited thick filter cakes in spite of very high eroding velocities. The muds having good filtration characteristics deposited thin filter cakes at equilibrium circulating velocities well within tile range of those in a drilling well. It was observed that filter cakes deposited during stagnant filtration were quite difficult to erode by mud circulation. The - rate of crosion computed from the rate of filtrate accumulation after equilibrium cake thickness had been reached was in reasonable agreement with the rate of erosion obtained by direct observation. Continuous mud circulation usually caused the permeability of the filter cake to decrease with time. INTRODUCTION Many of the difficulties encountered during tile drilling of a well have been attributed to the loss of water from the mud and the attendant deposition of solids upon the walls of the hole. Past experience has shown that a reduction of the filtration rate of the drilling fluid eliminates or greatly reduces these difficulties. Definite filtration requirements, however, are hard to establish for a given set of conditions. This is due. in part, to the fact that the usual filtration test performed upon mud doe? not simulate well conditions as closely as desirable. The filtration characteristics of a mud are customarily determined by means of the standard low-pressure API wall-building tester.' In this instrument a filter cake is deposited upon a horizontal bed under a pressure differential of 100 psi. The rnud is quiescent during the filtration period. In actual practice. mud filtration occurs within a well under quite different conditions. One of the major differences is that mud flows upward across the filter bed as the filter cake forms. This undoubtedly produces a change in the filter cake which cannot be reflected in the results of the API test. The laboratory work described in this paper had as its primary objective a better understanding of the influence of mud circulation upon the thickness and ,characteristics of the filter cakes deposited under conditions similar to those existing in a drilling well. ANALYSIS OF PROBLEM Once a permeable formation is penetrated by the bit, filtrate from the mud flows into the formation. 'he mud solids plaster against the walls of the hole, forming a filter cake. If the mud column is stagnant, that is, if it is not being circulated. the filter cake will increase in thickness until the hole is filled. Prior to the time that the hole is filled, the thickness of filter cake existing at any given time will be a function of the filtration characteristics of the mud, the temperature, and the pressure differential. The effects of these variables have been investigated in the past for both flat bed filtration2'3 and for radial filtration.' When the mud is circulated in a hole in which a filter cake i. being deposited. some of the solids that would ordinarily deposit in the filter cake will be carried away by the eroding action of the mud. This will limit. filter cake thickness. Some work has been done to determine the effect of flow upon the filtration rate in a circulating mud system' but little work has been done upon the factors which determine the filter cake thickness existing in a circulating system. On first sight it would appear that the major factors controlling filter cake formation in a circulating system should be: 1. The rate of deposition of solids from the mud. 2. The erosive force that the flowing mud exerts upon the filter cake. 'A. The erodabilitv of the filter cake. 4. Any change in filter cake characteristics attributable to the scouring action of the mud. The rate at which solids are deposited from the mud will be controlled to a large degree by the filtration characteristics of the mud, the pressure differential. the temperature under

Jan 1, 1952

By W. B. Nelligam, J. Tittman

Refinements in radiation logging techniques during recent years have involved increasing usage of scintillation ditectors. These detectors produce voltage pulses whose heights are related to the energies of the gamma rays which initiate them. Analysis of the gamma-ray spectrum, as indicated by the pulse heights, yields information about the chemical elements composing the formations surveyed. Refined scintillation counter techniques can furnish chemical information concerning earth formations in situ, from a study of the gamma-ray spectra emitted by the formation either naturally or as a result of neutron bombardment."' Accompanying the rising interest in gamma-ray scintillation spectroscopy, there has been increased activity in the development of accelerator-type neutron sources (in contrast to encapsulated chemical-mixture sources). Such neutron generators are attractive for several reasons: (1) they greatly reduce radiation hazards to personnel; (2) there is a great reduction in contamination danger if they are lost in the hole; (3) they can produce larger neutron intensities than can conveniently available encapsulated sources; and (4) they are capable of being pulsed, thus permitting new techniques in logging. In the past, both accelerator and encapsulated neutron sources have been used by others in conjunction with scintillation-detector pulse-height analysis. The results have not been too encouraging, due to the interference among different gamma-ray spectral lines and to the fact that the gamma-ray peaks were not too clearly distinguishable above the large and ill-defined background "noise".'.' This paper is a status report on laboratory studies of a technique using a borehole accelerator as a neutron source, which gives an improved scintillation spectrum, thus permitting more accurate chemical analyses of the formations penetrated. The Schlumberger-accelerator neutron source' is presented; the origins of inelastic and of thermal-neutron, capture gamma rays are dis- cussed, and results are given for some laboratory measurements performed in borehole geometry. THE NEUTRON GENERATOR-TUBE ACCELERATOR In the attempt to develop an accelerator neutron source which would have the desirable properties outlined in the "Introduction" and which would operate satisfactorily under borehole logging conditions, a small-diameter, cylindrical, neutron generator tube has been developed. This tube utilizes the principle of accelerating deuterons (heavy hydrogen nuclei) by a high voltage so that they bombard a tritium (heavy, heavy hydrogen) target. The resulting reactions produce large numbers of neutrons of 14 Mev energy. Furthermore, the tube is permanently sealed and, thus, the use of pumping techniques in the sonde is avoided. The tube consists of a pressure control, an ion source in which the deuterons are stripped of their electrons, an accelerating gap down which the deuterons are sped by the high voltage, a secondary electron suppressor and a tritium-loaded target. Obviously, the tube requires auxiliary circuitry in a sonde for the control of the ion source, tube pressure, high voltage and other operating variables. Neutron yields, both continuous and pulsed, have been produced under simulated field conditions in the range between 1 and 10 times those conventionally used in neutron logging applications. For the experiments discussed in this paper, neutron-pulse repetition rates in the range between 500 and 5,000 pulses/sec are adequate. Furthermore, by making suitable adjustments in operating conditions, one can vary the pulse width from a minimum of several microseconds up to dc. As will be seen later, the present application to well logging does not require that one have available pulses of neutrons which are appreciably shorter than the time it takes for fast neutrons to slow down to thermal energy in water. As a consequence, much of our interest has been directed towards the operating characteristics of the tube with pulses longer than about 10 microseconds in duration. Our experience to date shows that tubes can be operated in this manner with reasonably constant, average neutron outputs.