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By Roswell Johnson

To know the rate of decline of oil wells is very important, yet ordinarily we are prevented from getting this rate because the oil from several wells is put into one or a few tanks as soon as the wells have been connected to the power. The object of doing this is economy of tanks, pipes, labor and accounting. FIG. 1. But we are making a great sacrifice for this small economy. The advantages of the individual records are that we would know better: 1. When the well has ceased to be profitable. 2. When a well has gotten into bad condition. 3. The real record of the yield per acre and the rate of decline of the wells. The lease total is too mixed as to age of the wells for existing records to have much significance. To obtain such individual well records and more accurate logs are the great desiderata of satisfactory management of oil properties. 4. Which wells show the slowest decline-information which is valuable in locating new wells or purchasing adjoining leases.

Jan 6, 1917

Both drilling and production of oil decreased in Rumania during 1935. The drilling totaled 1,042,041 ft. and production 61,973,938 bhl., being a decrease of 8.5 and 1.6 per cent, respectively. Wells completed during the year totaled 198, of which 168 were oil wells, compared with 238 and 215, respectively, of 1934. No important discovery was made during the year. The only new field found was Valea Calugareasca, where the Credit Minier completed a small well on the westward plunge of the anticline. Drilling higher on the structure will probably prove the field to be of greater importance than the results of the discovery well indicate. The only other developments of importance were the extension of the Bucsani field, the opening up of a new area on the northeast of the salt at Moreni by Credit Minier and a minor extension of the West Gura Ocnitei field. At Calinesti, on the cast extension of the Moreni field, exploration work continued and a gas well was completed, but oil in commercial quantities is yet to be found.

Jan 1, 1936

By Jerry A. Selvaggi, Richard B. Jacob

The High Gradient Magnetic Separator (HGMS), also known as the High Gradient Magnetic Filter (HGMF), is an effective device utilized in magnetic separation. This method of separation is gaining popularity in the separation of submicron material containing feebly magnetics. The most common use of this type of separator is in the removal of iron stained particles to improve the brightness of kaolin clay. Potential areas for use of this magnet include: (1) desulfurization of coal (2) the clean-up of mill effluent (3) removal of fine particulate matter from municipal and indus¬trial waste (4) desulfurization of oil Design, operation, and applications of this separator are further dis¬cussed in papers by R. R. Oder (1975), and R. R. Oder and B. I. Horst (1975). This device is a system composed of an iron enclosure of high per¬mability material housing a set of electrical coils, a canister en¬closing a magnetic matrix (stainless steel wool), an electric power supply and a hydraulic system to maintain specified operating tempera¬ture conditions. See Figure 1.

Jan 1, 1984

By Rupert Spivey

Kennecott Copper's Chino mine is one of the oldest copper mines in the United States, having been operated by the Spaniards as early as 1800. At that time haulage methods were indeed primitive. Indians brought the ore up on "chicken ladders" in leather buckets. It was then melted into crude bars and hauled to Chihuahua by mule train. There are still remnants of the old trails near the present pit where the mule hooves cut grooves in solid rock. The mines were worked as underground operations off' and on until open-pit mining was started in 1910 by John Sully and Daniel C. Jackling. Since that time many haulage methods have been used: mule carts, steam and electric trains and finally large haulage trucks, now believed to be the best method for this mine. The general shape of the pit is oval with dimensions of 1% miles by 1% miles. Working benches are 50 ft high, with the bottom bench at 5750 ft and the top at 6450 ft above sea level.

Jan 1, 1964

By Francis T. Havard

A Reply to the Discussion by Mr. Francis T. Havard of the Paper by Mr. Henry W. Edwards, which was read at the Atlantic City Meeting, February, 1.904. (Annual Meeting, February, 1905.) MR. HENRY W. EDWARDS, Grand Junction, Colo. (communication to the Secretary) :* Effect of Heat. It should be remembered that the coefficient of expansion of concrete is nearly that of iron; therefore in constructing flues the design must take account of this important detail. The cracking of flues, mentioned by Mr. Havard as having occurred at a temperature of 100° C. and above, . were possibly caused by neglecting to appreciate the extent of the expansion and contraction of the concrete. Effect of Atmospheric Moisture.-The moisture present in the atmosphere is exceedingly injurious to flues of any material, and those made of concrete are no more susceptible than of brick. My experience with flues of the Monier construction, having walls from 1.5 to 2.5 in. thick plastered upon a net-work of iron rods, was unsatisfactory from the standpoint of dura¬bility.

Mar 1, 1905

Minnesota School of Mines Experiment Station, University of Minnesota, Minneapolis, Minn For copies of Bulletins or a list of publications, address The Director Bulletins available are Bulletin 1, Iron mining in Minnesota, by C. E Van Barneveld (1912), $1, 2, Concentration tests on Mesabi ores, by W R Appleby and E Newton (1913), 3, Preliminary concentration tests on Cuyuna ores, by W. R Appleby and E Newton (1915), 4, Bibliography of Minnesota mining and geology, by W. Gregory (1915); 5, Manganiferous iron ores of the Cuyuna district, by E Newton (1918), 7, The future of the Lake Superior district as iron-ore producer, by E W. Davis (1920), 9, Magnetic concentra¬tion of iron ore, by E. W Davis (1921); 10, Ball mill crushing iii closed circuit with screens, by E W Davis (1925); 12, Minnesota manganiferous iron ore in relation to the iron and steel industry, by T L Joseph and S P Kinney (1927). A mining directory of Minnesota, by J J Craig, is published annually.

Jan 1, 1933

Geological Survey of Alabama, University, Ala. A complete list of publications will be sent upon request Requests for publications should be addressed to Walter B. Jones, State Geologist, University, Alabama Publications of the Survey, of most interest to the general public, comprise (1) Special Reports, (2) Monographs, (3) Bulletins, (4) Circulars, (5) Maps The Special Reports mostly deal with a special coal field, mountain, valley, or series of formations, while the Monographs treat of some industry, resource, or group of organisms throughout the state or as far as it extends The Bulletins deal with a variety of subjects, but are usually smaller and less technical than the Monographs and Special Reports Maps now available comprise a map of the Cahaba coal field, one of the Coosa coal field and a small geological map of the state, in addition to a large geological map, 28 by 42 inches, 8 miles to the inch (1926) (also bound with Special Report 14) Price of the large geological map of Alabama, folded in manila envelope, is $1.03 postpaid

Jan 1, 1933

By E. E. ELLIS

THE zinc deposits under development by the Universal Exploration Co., a subsidiary of the United States Steel Corporation, are close to Jefferson City, Tenn., a small town about 30 miles northeast of Knoxville, Tenn., and are served by the Southern Railway. The first zinc mining in this district was in the old Mossy Creek mine at the north edge of the town, opened for oxidized ores in 1871 and later worked for sulfide ores. An oxidized zinc ore deposit (carbonates and silicates) was opened near New Market in 1900 and worked through the war period by the G1'8sselli Chemical Co., producing around 25,000 tons of oxidized ore. In recent years a sulfide deposit has been found and developed but is not producing. About 1% miles south of this a small body of sulfide ore was developed by open cut, but the operation was unsuccessful. The Universal Exploration Co. entered the district in 1926 and, after extensive exploration, has been operating an oxidized ore-body since Janual:y, 1928, and has been developing a sulfide ore body since July, 1928.

Jan 1, 1929

By George E. Beavers

It is well to understand at once that converter gas is always mixed with copper blast-furnace gas at Copperhill before entering the Glover tower, and, at times, roaster gas is also added along with the converter and blast-furnace gas. Fig. 1 is introduced to show the variations in the sulfur dioxide content of the gas and the ability of the acid-plant operator to control the variations, resulting in a very even grade at the Gay-Lussac tower entrance. The sulfur dioxide content of the gas entering the Gay-Lussac tower must be low and regular to assure low niter consumption. The blast-furnace gas alone is rather uniform as regards sulfur dioxide content and the variations, as shown in Fig. 1, are chiefly caused by converter changes—out for charging, slag pouring, copper pouring, and, of course, the "finish." The combined gas during the converter finish is, at times, almost free of oxygen and the sulfur dioxide content reaches 13.0 per cent. At such times air is added at the Glover tower.

Jan 1, 1934

Reports on process control computer projects at the Kamioka Plant are summarized from the Japanese technical literature (see references). Fluidized Bed Roaster and Sulfuric Acid plant1 A small process control computer (Yokogawa CCS-200 with an OKITAC 5090P 8K central processor) was installed in the Kamioka smelter in the summer of 1964. A year-and-one-half installation project followed, which successfully brought the roaster and acid converters under computer control. The process plant consisted of two 130 RPD fluosolids roasters and two Monsan to acid units. The roasters are charged with lead-zinc concentrates with SO2 from the roaster processed to sulfuric acid. The computer system provides the following functions: 1) log data from the plant--the input is taken from 50 analog signals, 130 high-low digital alarm points, and 50 analog high- low alarm points; 2) set-point control on 18 control loops; 3) direct digital control (DDC) on eight control loops for which conventional controllers were difficult to adapt due to abnormal loop characteristics or delay times; 4) execution of a safe shut-down procedure

Jan 1, 1969

By H. R. Hanley

ELECTROLYTIC zinc produced from sulfate solution and, with pure lead anodes is always contaminated with a small and varying percentage of lead. The purpose of this investigation is to determine the characteristics of several lead alloys and their influence on the products of electrolysis and power consumption. The study of alloyed anodes has been a matter of interest for some years. Tainton, Taylor and Ehrlinger prepared a highly interesting paper1 on this subject a year ago, in which the effects of silver-lead and some ternary alloys of lead-silver-tin were described. The present paper describes a study of various combinations of 10 metals. The general plan of the work provides for the fabrication of the alloyed anode, its adequate conditioning by preparatory electrolytic operation, the production of cathode zinc and its analysis, the determination of the anode polarization, and other features. A list of the anodes tested, showing composition and polarization at various current densities together with the lead assay of the cathode zinc, is shown in Table 1.

Jan 1, 1930

By M. M. Wong, D. E. Couch, G. M. Martinez

The Bureau of Mines electrowon hafnium metal with an average oxygen content of' 150 ppm at 700°C from an electrolyte containing 27 wt pct LiCl, 62 wt pct RbCl, and 11 wt pct HfC14. The average anode and cathode current efficiencies were 90 pct at anode and initial cathode current densities of 86 amp per sq ft. Haf-nium metal with an average oxygen content of 440 ppm was electrowon at 800oC from an electrolyte containing 90 wt pct KC1 and 10 wt pct HfCl4. The average anode and cathode current efficiencies were similar to those obtained in the LiCL-RbCl-HfCl, electrolyte. The chlorine gas given off at the graphite anode was vented through either a silica or a graphite tube to prevent cell corrosion. THE current method for the commercial production of high-purity hafnium is the thermal decomposition of Hfl4.1 The iodide method is not adaptable to continuous process techniques. Nettle, Hiegel, and Baker2 studied the electrorefining of hafnium from hafnium sponge containing 800 ppm oxygen. They failed to obtain hafnium with 600 ppm oxygen in their initial deposits and obtained AEC specification for oxygen only after 75 pct of the soluble hafnium had been removed from the electrolyte. Calculations using their data indicated this was approximately 4 lb of hafnium. The electrolyte was then used to produce approximately 3 lb of hafnium with a low oxygen content. However, no data are shown concerning the amount of anode material initially used or what percent of it was dissolved, therefore, results are not suitable for evaluation of a continuous operation. In general, it was not possible to consistently obtain low oxygen content metal with the electrolytes described by Nettle, Hiegel, and Baker. Wong, Hiegel, and Martinez3 investigated the electrorefining process for hafnium and showed that even by strict control of electrolyte composition only relatively low oxygen reduction could be obtained. The oxygen contained in the hafnium anode material tended to transfer to the cathode deposit and only a limited purification was possible. Both the "iodide" and the "electrorefining" processes depend upon hafnium sponge as a starting material. The sponge is normally produced by magnesium reduction of HfC14 ' and does not meet AEC specifications for hafnium metal. Since only 30 pct of the anode feed could be utilized3 in the electrorefining cells, the Bureau of Mines developed an electrowinning process. HfC14 was used as the feed material for the electro-winning process described in this report. Many of the electrolytes used in the electrorefining studies3 ap- peared to be suitable carrier-electrolytes for HfC14. However, in the initial studies on electrowinning, it was desirable to use electrolytes that had low solidus temperatures and could be operated over a wide temperature range to investigate parameters of the process. Therefore, electrolytes containing LiC1, NaC1, KC1, RbC1, CsC1, and HfC14, in various combinations were explored. EQUIPMENT Chlorinator. Hafnium carbide was chlorinated to produce HfC14 in the batch-type chlorination shown in Fig. 1. Chlorination temperatures were measured with a thermocouple placed in the center of the HfC charge. A flow meter was used to monitor the helium and chlorine. The exhaust side of the silica chlorina-tor tube was equipped with a flask for collecting organic material released during the initial heating of the HfC. The temperature of an internal heater, which extended from the HfC14 condensing flask to the hot end of the chlorinator, was adjusted to prevent the HfC14 from condensing before entering the collection flask. Helium and excess chlorine were exhausted through the lid of the collection flask to an aqueous NaOH solution. Sublimer. Initial studies were conducted using a sublimer, Fig. 2, made by placing a 13/8-in. OD nickel thimble 11 in. long, inside a 11/2-in. ID nickel bell 12 in. long, and locking it in place. This unit was loaded with HfC14 and partially immersed in the molten electrolyte for sublimation directly into the electrolyte. In another sublimer shown in Fig. 3, the HfC14 was contained in a "resin reaction flask". Quartz wool, previously heated to 600aC, secured between two nickel wire screens, was placed just above the HfC14 powder. The lid contained a vacuum outlet, a gage, an argon inlet, and an air-cooled pipe for condensing the HfC14. This sublimer was evacuated and heated. The sublimation temperature was not critical and the sublimer operated satisfactorily at all temperatures between 250" and 350°C. Electrolytic Cell. The electrolyte chamber, Fig. 4, was made of mild steel 8-in. schedule 20 pipe, 30 in. long. The exterior was metallized with a Ni-Cr alloy. The electrolytes were contained in a 16 gage nickel or iron liner with a nickel heat shield on top. The cell was heated by a resistance furnace. A 21/2-in. ID by 25 in. long air lock was connected to one port of a two-port cell cover assembly through a slide valve. The cover assembly of the air lock was electrically insulated from the cell and was equipped with a rubber sleeve that provided for the passage of the cathode lead. This allowed the cathode deposits to be removed and a new nickel cathode to be introduced without allowing air to enter the cell. A tube-rod assembly was bolted to the other port on the cell cover assembly and was sealed by a packing seal. The tube-rod assembly consists of a graphite

Jan 1, 1970

By C. T. Liu, J. Gurland

The deformation behavior in tension of spheroidized carbon steels was studied at room temperature as a function of carbon content, 0.065 to 1.46 wt Pct, and carbide particle size, 0.88 to 2.77 p. It was found that the Hall-Petch strength-grain size relation is directly applicable to the yield and flow stresses of the two lower-carbon steels , 0.065 and 0.30 pct C. The strength data for the medium- and high-carbon steels, 0.55 to 1.46 pct C, also satisfied the Hall-Petch relation, provided that these data are based upon the particle spacing. Beyond 4 pct strain, the flow stress data of all the steels studied could be represented by the same Hall-Petch relation with dinerent spacings for grain boundary and particle strengthening. The behavior of the higher-carbon steels was consistent with the postulated formation of a dislocation cell network during processing and initial deformation (up to 4 pct strain). The cell size was assumed to be equal to the planar particle spacing. The true stress at the ultimate tensile strength was also found to be a function of the particle spacing. At a given temperature and strain rate, the yield and flow stresses of carbon steels depend on the type and dimensions of the microstructure. Starting with the work of Gensamer et al. in 1942,' experimental studies on pearlitic and spheroidized carbon steels revealed that the strength of steels is a function of two main parameters: the ferrite grain size2'3 and the carbide particle spacing;1'4'5 on this basis, two different strengthening mechanisms have been developed to apply to steels of low and high carbon contents, respectively. In polycrystalline iron and mild steels the grain boundaries are regarded as the major structural barriers to slip. The relation between strength and grain size is generally represented by the Hall-Petch equation which is based on a linear proportionality between strength and the inverse square root of the average grain size.2'3y677 However, Gensamer et al.' and Roberts et related the yield strength of medium -and high-carbon steels to the carbide particle spacing alone, and they found a linear relation between the logarithm of the mean free path in the ferrite and the yield strength in both spheroidized and pearlitic steels. By means of the electron microscope, Turkalo and LOW' extended the study to finer structures; they concluded that the logarithmic relation is not valid for the entire range of microstructures unless grain boundaries are also included in the measurement of the mean free path. For the specific case of spheroidized steels, Ansell and aenel' found that the yield strength data,4'5 when plotted as a function of mean free path, fit the Hall-Petch equation; however, T'ysong found that the same data fit the 0rowanl0 relation if a planar inter-particle spacing is used. Recently Kossowsky and ~rown" studied the strength of prestrained spheroidized steels, 0.48 and 0.95 pct C, and concluded that the strength due to the carbide dispersions varies linearly with the reciprocal of the square root of the mean free path between carbide particles and dislocation networks. Such networks were first observed by Turkalo." The conclusion common to all these studies is that the available slip distance in the ferrite is the most important variable in determining strendh. Previous work on carbon steels is restricted to limited composition and strain ranges. The mechanism which governs the flow properties is not clearly understood, and, in particular, little is known about the composition dependence of the transition between grain boundary strengthening and particle hardening. The purpose of the present work is to investigate the strengthening mechanism in spheroidized steels over a wide range of carbon content, 0.065 to 1.46 wt pct, and plastic strain, yielding to necking. The spheroidized structure was chosen because of its relative simplicity and the relative ease of control and measurement of the structural parameters. The experimental work is limited to tensile testing at room temperature at constant extension rate. The effects of the carbide particles on the fracture behavior of spheroidized steels are discussed elsewhere.13 EXPERIMENTAL PROCEDURE Eight different grades of vacuum-cast carbon steels were supplied in the form of forged and rolled plate by the Applied Research Laboratory of the U.S. Steel Corp. The compositions furnished with these steels are given in Table I; the carbon content ranges from 0.065 to 1.46 wt pct, or from 1.0 to 22.3 vol pct of carbide. The steel plates were cut transversely into rods a little larger than the test specimens, 1 in. gage length, i in. diam. The rods were austenitized in air (enriched with CO by a consumable carbon-rich muffle) at 50° C above theA, orA., temperature for 2 hr and then quenched in oil with vigorous stirring. The as-quenched rods were tempered in two stages in order to obtain the desired distributions and sizes of carbide particles. The rods were first tempered at 460° C for 10 hr and then at 700" C for periods ranging from 4 hr to 3 days, in vacuum. After final machining, all specimens were vacuum-annealed again at 650°C for 1 hr in order to relieve residual stresses. The tension tests were carried out in two steps. The initial part of the load-strain curve, up to about 2 pct strain, was determined on a Riehle testing machine with an extensometer of small strain range, 4 pct strain, in order to obtain the yield and initial flow piopertiesi As soon as the first part of the test was finished, the specimen was placed in an Instron testing machine equipped with a strain gage extensometer with a maximum strain range of 50 pct. The load-strain curve to fracture was

Jan 1, 1969

By Garth M. Crosby, F. McIntosh Galbraith, Bronson Stringham

THE Hercules mine is located in the northeastern section of the Coeur d'Alene district, approximately 1 1/2 miles north of the town of Burke, Idaho. Surface indications of the ore deposit were first discovered in 1886, but regular mine production was not started until 1902 and was continuous until April 1925, when the known ore had been extracted. Incomplete records show that from 1912 until operations were suspended the mine produced 2 1/2 million tons of ore containing 9.4 pct lead and 7.7 oz of silver per ton, together with an estimated 2 pct zinc, 0.3 pct copper, and 20 pct iron. This operation was the first in. a series of mining enterprises culminating in October 1947 with the consolidation of Day Mines, Inc. In the same year it was decided to unwater the levels below the collar of the Hercules shaft in the hope of finding some indication of a recurrence of ore. The unwatering operation has been described in a. previous paper.' The initial exploration, following recapture of the workings, showed sufficient promise to warrant a detailed study of the mineralogy with modern techniques. The general geo1ogy of the Coeur d'Alene district, including a detailed description of the rock types encountered, has been comprehensively treated by Ransome and Calkins' in their classic paper, and only local background description, therefore, is felt to be appropriate here. The Hercules deposit transects a portion of the trough of a broad south-trending synclinorium which has been greatly complicated by faulting. More locally, it lies within a block of ground bounded on the east; by the O'Neil Gulch fault, a steep north-south overthrust of considerable magnitude, and on the west by a monzonite stock, the outcrop of which is 1/2 mile or more wide and 5 miles long. The country rock is composed of thin to medium-bedded argillites and argillaceous quartz-ites of the Prichard and Burke formations, the oldest members of the Pre-cambrian Belt Series of sediments in the area, believed to be of Algonkian age. The contact between them is a conformable gradation. The argillite is colored gray to tannish-gray and is fine-grained, compact, and generally massive in structure. Under the microscope the unaltered argillite is seen to be composed principally of anhedral quartz and a few feldspar grains which were at one time presumably partly rounded sand grains, but as a result of recrystallization and cementation by silica, the interstices are now almost obliterated and quartz grains show crenulate boundaries. The sizes of these crystals vary from 0.5 mm down to 0.1 mm in greatest dimension. In all specimens sericite comprises 10 to 20 pct of the rock and is present abundantly between most of the grains as flakes or shreds which vary considerably in size. Sometimes they form a fine felt-like mat or aggregate, and sometimes flakes are seen which appear to be good muscovite. In some specimens, separated rhombic-shaped carbonate grains are abundant, and in some instances these have been changed to sericite. Mining operations to date have explored the Hercules vein to a maximum vertical depth of 3600 ft below its outcrop, and along a maximum strike-length of 3600 ft on certain of the lower mine levels. The main orebody is irregular in outline, extending over a variable strike-length of 400 to 1500 ft; and it is intersected by a strong transverse fault that has been traced from the surface to the bottom level. This has been named the Hercules fault, and apart from the vein itself, it is the most prominent structural feature in the mine. There is good evidence that it existed prior to the introduction of ore solutions and may have influenced ore deposition, but it was also the locus of important post-ore displacement and shows a progressive right-handed horizontal component reaching 200 ft on the deeper levels. Its vertical component is not definitely known but may be considerably greater. The fault strikes 20° N to 50° E and dips westerly at angles of 70" to 45", flattening in dip where it crosses the original orebody from east to west between 1000 and 1600 ft below the surface. At about 3000 ft in depth the Hercules fault is joined by a vertical fault of similar strike, and the major post-ore dis-placement below their junction is taken up along this vertical branch of the structure, now called the Mercury fault. Recent work has been concentrated in this vicinity. Another structural feature of special geologic interest, though of little economic importance, is the occurrence of a porphyritic dike in this area. This lies a short distance above the Hercules fault, essentially parallel to it, and is 5 to 15 ft in thickness. It appears at first glance to cut the mineralization, suggesting push-apart relationship, but small stringers of the vein minerals have been observed to penetrate the dike for a matter of inches at several points. The dike is thought to be related to the monzonite intrusion. A vertical longitudinal projection of the mine is shown in Fig. 1, which illustrates most of the features discussed above. The Hercules vein was deposited along the course of a strong, persistent shear zone that now appears as a braided network of gouge seams running through more or less crushed and shattered country rock. It strikes 70° N to 80° W and dips southerly at an average of 75". Barren parts of the structure vary in width from less than 1 ft to more than 15 ft. The width of mineralized segments may be double that. Although the evidence is not conclusive, pre-mineral, normal movement along the zone may be 1000 or 1500 ft. The horizontal component is unknown. Post-ore movement appears to have been

Jan 1, 1954

By Joseph H. Allen

In the old Kentucky reports, made by the Survey of which Prof. Owen was director, the veins of the Western Kentucky coal-field were numbered from 1 up to 12. Later Prof. Shaler used letters, beginning at the top. For instance, Owen's 12 is Shaler's A.

Jan 1, 1888

By R. E. Zimmerman

Wide attention is being placed upon various methods for cleaning the fine sizes of bituminous coals. The author describes and analyzes the results achieved on wet concentrating tables of modern design on unclassified feeds of --1/2 in. particle sizes, including tests on high-sulphur coals.

Jan 9, 1950

By H. H. Podgurski, J. C. M. Li, Y. T. Chou, F. N. Davis, R. A. Oriani

When an Fe-2 pct A1 alloy is nitrogemted at 500ºC with a gus tnixture (NH3-H2) in which the nitrogen activity has been kept Lou] enough to avoid the formation of iron nitride, a two-phase alloy is generuled which consists of AlN particles and a ferrite phase cotaining a heavy network of dislocations. The amount of nitrogen contained in such an alloy, when equilibrated with the nitrogenating atmosphere, far exceeds both that needed to satisfy the normal solution requirements of a Fe and that needed to convert all of the aluninuwi to AlN. This excess nitrogen is accounted for as being trapped on dislocations, adsorbed at the ferrite -AlN interface, and as an en-Iuznced lattice solubility in strained ferrite. This excessive uptake of nitrogen had previously been attribuled by other investigators to the formation of a nonstoichiornetric aluminum nitride. Isotope exchange experiments revealed various amomts of exchargeable N14 present in the originally nitrided samples that could not be removed by reduction with HS at 500ºC. This exchangeable nitrogen has been identified us that bound to the AlN -ferrite interface. Estimates of inter facial areas in alloys containing -3 pct by weight of A1N are as high as 10 sq m per g of alloy. ThE first1 of this series of papers described the experiments forming the basis for the elucidation of the mechanism of the formation of aluminum nitride particles within an Fe-A1 alloy. It was found that not only are dislocations necessary for the nucleation of the AlN particles but also the nitriding reaction in turn produces a dense network of dislocations in the ferrite matrix. It was also observed that nitrogen in excess of that needed for the formation of stoichiometric AIN is taken up by the alloy without the formation of an iron nitride. The present paper is an analysis of the excess nitrogen sorbed by the nitrided Fe-A1 alloy which considers the structure generated by the formation of the aluminum nitride particles. Because of the high density of dislocations, this system has proved to be quite useful in studying nitrogen-dislocation interactions. Wriedt and Darken2 have already reported such studies in a cold-worked ferritic steel. EXPERIMENTAL Nitrogen Sorption. The specimens of this alloy were cold worked (50 pct reduction) to 0.011 in. thickness and chemically cleaned in a 2:1 concentrated phosphoric acid-50 pct hydrogen peroxide solution before nitrogenation. The flowing nitrogenating atmosphere (11 pct NH3-89 pct H2) was established before the ni- trogenating temperature was reached. The same gravimetric and gas-flow equipment described in an earlier paper1 was also used in this investigation. Changes in nitrogen concentration were followed gravimetrically when establishing the isotherms. In the isotope exchange experiments, concentration changes were followed volumetrically. In some instances chemical analyses (Kjeldahl method) were used to check for material balances in both the gravimetric and volumetric procedures. Our objective was to obtain reversible nitrogen sorption isotherms for alloys equilibrated with NH3-H2 gas mixtures over a large range of nitrogen activity* and temperature. The *Defined as equal to PN H /P3/2 ,where P corresponds to partial pressure in atmospheres. Actually the nitrogen activity, aN, in the alloy equals K PNH3 /p3/2H2, where K is the equilibrium constant for the reaction NH3 = N + 3/2H2. upper limit for nitrogen activity was below that which would produce iron nitride; for reasons which will become apparent later, most of the sorption studies were made in the temperature range between 400" and 500'C. The alloy sample studied most extensively in this investigation was given a series of successive reductions (100 pct H2) and nitrogenating treatments at 500°C to attain a stabilized structure. Presumably some dislocations were lost during these treatments. Throughout most of the sorption studies, temperature was held at ± 1°C and the gas-phase composition was held to k0.2 pet NH3. Based upon the results of numerous diffusion experiments with this alloy* the 'Results to be published in a third paper of this series. times chosen for equilibration were considered more than adequate, Nitrogen Isotope Exchange. The first exchange experiments were carried out by circulating a measured quantity of H2 and NH3* (with a predetermined isotopic *The NH3 was synthesized over a synthetic ammonia iron catalyst using H2, and N2, enriched with N15. AS the NH3 was formed it was removed continuously from the gas phase by circulating the mixture through a refrigerated (78ºK) trap. When most of the nitrogen had been converted to NH3, it was distilled from the trap into a glass storage vessel. composition, N15/N14) over a nitrided specimen at 450°C in a closed glass system. Attempts to reach the exchange limit isothermally, i.e., an isotopic ratio (N15/N14) in the gas phase identical with the extractable nitrogen in the nitrided alloy, were futile. The exchange rate between the gas phase and the alloy was too slow, involving many days. To circumvent the slow exchange with the gas phase, measured amounts of N15 and N14 were introduced into the alloy by nitrogenating with an NH3-H2, mixture at 500°C containing an N15/N14 ratio of 7.76; the charged specimens were then sealed by allowing an oxide film to form over them in air, and in this final condition the specimens were subjected to 500°C in vacuum for various times during which isotope exchange was allowed to proceed within the specimen. It was estab-

Jan 1, 1970

By Rudolph G. Wuerker

THE scarcity of values of tensile strength of rocks has been explained by the lack of successful testing procedures. In the case of mine rock a description is given' of the difficulties encountered in testing a cylindrical specimen, such as a core, by conventional methods. Over a period of years the following method has given definite and reproducible results with the weakest as well as with the strongest rocks. It does not completely supersede the use of cores with special fixtures but is a supplement in all cases where cores cannot be obtained, as from soft rocks, or in cases where it is less expensive to prepare test specimens by cutting them out of the rock instead of drilling cores. Principle and equipment are the same as for the test for tensile strength of hydraulic-cement mortar.' The test specimen, Fig. 1, has the shape of a briquet. While in the original cement mortar test the briquet is cast in a special mold, it is prepared from rocks in different ways, depending on how easily they can be cut and shaped. Soft rocks, which cannot be core-drilled with a carboloy or diamond bit, are simply hand-cut. Only two dimensions need be watched. The first is the 1-in. diam at the narrowest cross-section of the briquet. The other critical measure is the radius of curvature of the waistline, as the roller supports in the grips have a fixed distance. This radius is ground out of the solid by means of a carborundum grinding wheel having a 3/4-in. radius. Medium hard rocks can be core-drilled with a carboloy bit. The resulting core can be used for nondestructive sonic testing first, and after that for any destructive test. By using an EX-bit and by carefully placing the coreholes, preferably by using a tenplate such as shown in Fig. 1, it is possible to obtain from the rock a punched sample from which numerous tensile briquets can be made. The outside radius of the EX-bit differs from the radius of curvature of the briquet by 1/8 in., but this still permits placing and aligning the specimen in the grips. In the case of bedded rocks the core might have bedding planes normal to the plane of the briquets, and rocks can be tested in any arrangement of the bedding planes desired. Hard rocks, limestones, igneous, and metamor-whic rocks can only be diamond-drilled or diamond-cut. Here the method of getting the tension briquets by accurate placing of EX-drill holes is especially economical. The tops of the briquets made from hard rocks cannot be rounded; they are straight cuts made with a diamond cut-off saw and rounded off on a polishing wheel. Results: As long as specimens broke over the waistline the results were considered acceptable. Further statistical treatment of the tests' showed a satisfactory percentage of standard deviation. The tensile strength values obtained by this method do not represent true values because of the stress concentration caused by the curvature of the side of the piece and because of the closeness of the grips. The ratio of maximum to average stress at the plane of failure has been determined to be about 1.75." All tensile strength values listed in Table I are corrected accordingly. To avoid this stress-concentration, if there are a sufficient number of cores, tensile strength can be measured by imbedding the cores in mortar in the two outer briquets in the gangmold.4 Strain-Measurements: The applicability of the briquet specimens for strain observations was tested in the case of sandstone and shale. Two element Rosette SR-4 strain gages were used. Young's modulus and Poisson's ratio, both in tension, were computed and found to be different from those in compression, determined during the same test series and from the same rock, see Table I. References 1 L. Obert, S. L. Windes. and W. I. Duvall: Standardized Tests for Determining the Physical Properties of Mine Rock. U. S. Bur. Mines R.I. 3891 (1946). 2 Test for Tensile Strength of Hydraulic-Cement Mortar, ASTM Standard C 190-44. S F. O. Auderegg, R. Weller, and B. Fried: Tension Specimen Shape and Apparent Strength. Proc. ASTM 11939) 39, pp. 1261-1269. 4 API Code 32.

Jan 1, 1956

By A. E. Back, S. F. Ravitz

When manganese ores are leached with sulphur dioxide, a large part of any zinc in the ore usually is extracted with the manganese.1 In the dithionate process,2,3 in which the manganese ore is leached with dilute sulphur dioxide gas and the manganese is subsequently precipitated by adding lime to the pregnant solution, the zinc is precipitated with the manganese. These facts suggested the possibility of applying the dithionate process to the recovery of zinc from oxidized ores or calcined sulphide concentrates. Accordingly, an investigation was begun, and preliminary tests were made on several ores and concentrates. The results were encouraging, but the need to devote available funds and personnel to other projects made it necessary to drop the work before the various factors involved in the process could be studied in detail. A brief mention of the process,4 however, elicited a number of inquiries, so that it appeared desirable to report the results of the preliminary work. The dithionate process as applied to zinc may be summarized briefly as follows: The ore or calcined concentrate is leached with dilute SO2 gas to dissolve the zinc, largely as zinc sulphate, in accordance with the following equation in which [ZnO] represents the "oxide zinc" in the ore or calcine: [ZnO] + SO2(g) + 1/2 O2 = Zn++SO4— [1] If, in the leaching step, the ore or calcine is suspended in calcium dithionate solution, the sulphate ion, as it is formed, is precipitated immediately as calcium sulphate, which is subsequently filtered off with the insoluble ore residue: Zn++SO4— + Ca++S2O6— = CaSO4(s) + Zn++S2O6- [2] An alternative procedure is to suspend the ore in water, filter off the insoluble ore residue after leaching, add calcium dithionate to the resulting zinc sulphate solution, and then filter off the calcium sulphate precipitate. In either event the zinc is obtained in the form of a pregnant solution of zinc dithionate, from which it can be precipitated as relatively pure zinc hydroxide by adding milk of lime: Zn++S2O6— + Ca(OH), = Zn(OH)2(s) + Ca++S2O6- [3] The zinc hydroxide precipitate is calcined to form the final zinc oxide product: Zn(OH)2 = ZnO(s) + H2O(g) [4] Theoretically the process is completely cyclic with respect to dithionate ion, the calcium dithionate required for reaction 2 being replaced by that formed in reaction 3. In practice, of course, small amounts of dithionate would be lost in the various filter cakes, depending on the economic washing limits. In the leaching step, however, some of the SO2 is oxidized to dithionale instead of sulphate, accord- ing to the following overall reaction: ZnO + 2SO2(g) + 1/2 O2(g) = Zn++S2O6-, [5] so that the dithionate losses are made up, at least in part. Ores and Concentrates Tested The analyses of the ores and concentrates tested are summarized in Table 1, which also includes the analyses of the calcines produced from the concentrates. The concentrates were roasted in a laboratory muffle furnace at the temperatures indicated until evolution of SO2 could no longer be detected; shallow clay dishes ordinarily were used; but calcines 2a and 2c were prepared in iron dishes, with the result that these two calcines were badly contaminated with iron scale. The "complex" concentrate, sample 1, consisted essentially of sphalerite grains containing inclusions of pyrite and other sulphide minerals too minute to be liberated by grinding within economic limits. In the " marmatitic" concentrate, sample 2, the iron was present largely as marmatite (that is, as iron sulphide in solid solution in zinc sulphide) rather than as pyrite inclusions. The zinc was present largely as the carbonate in the two "oxidized" ores, samples 3 and 4, and as the silicate in the "willemite" ore, sample 5. Leaching Tests The leaching cell was a 3-liter glass beaker placed in an electrically heated water bath manually controlled to maintain a temperature of 35°C in the leach slurry. The slurry was agitated by means of a flat paddle-type stirrer

Jan 1, 1950

By Frank C. Kelton

It is perhaps not widely realized that extraction and saturation processes carried out on oil well core samples alter the properties of these samples to varying degrees. On the other hand it is felt by some that quick-freezing of core samples increases their permeability and porosity significantly. Accordingly, laboratory tests were carried out on 49 pairs of horizontally adjacent samples in order to differentiate between the effect of quick-freezing per se on permeability and porosity of the samples, as distinguished from the effect of the identical saturation treatment on permeability and porosity of the companion samples. Also, additional field data were obtained on comparison of frozen vs unfrozen companion samples. LABORATORY INVESTIGATION OF FREEZING us SATURATION EFFECTS Procedure The samples used in these tests were two-cm cubes cut in horizontally adjacent pairs from cores from eight Gulf Coast and Mid-Continent wells, which cores had not previously been frozen. These samples were extracted with carbon tetrachloride, dried, and air permeabilities run in the conventional manner. They were then evacuated and saturated with brine of 25,000 ppm sodium chloride content, and porosities determined by gain in weight. The samples were partially desaturated by evaporation down to an average brine saturation of 68 per cent. One sample from each pair was quick-frozen by covering with dry ice after wrapping in a single layer of paper, and allowed to remain frozen for about two hours; the companion sample from each pair was not frozen. After thawing the frozen sample, all samples were immersed in tap water overnight in order to leach out most of the brine. Air permeabilities were re-run, and the samples were again saturated with brine to determine a second porosity value. For purposes of averaging of data, the samples were grouped according to four permeability ranges, from 0 to 10, 10 to 100, 100 to 1,000, and 1,000 to 3,840 md. Average permeability and porosity changes for the frozen vs the unfrozen adjacent samples are shown in Table 1. Discussion As may be seen from Table 1, the averages of the per cent permeability increases for the quick-frozen samples ranged from 3.8 to 12.9 per cent among the four permeability groups. The average changes among the four groups of unfrozen companion samples ranged from a decrease of 0.2 per cent to an increase of 9.3 per cent. There was no particular correlation of these changes with magnitude of permeability; however, the increase for each group of frozen samples paralleled the increase for the corresponding unfrozen samples. The differences between the two sets of values are believed to be a valid indication of the effect of the quick-freezing in itself, since the treatment of the two samples in each pair was identical except for freezing. The permeability changes which are strictly the result of the quick-freezing are shown in the sixth column of Table 1. These range from a decrease of 0.9 per cent to an increase of 4.0 per cent; the overall weighted average is 1.2 per cent, as compared to an average increase of 6.8 per cent caused by the saturation treatment of the samples not frozen. The average porosity changes are in general smaller than the changes in permeability, and range from a decrease of 2.3 per cent to an increase of 3.3 per cent. The overall weighted average change ascribed to the quick-freezing is 1.0 per cent of porosity. Many factors can contribute to the changes in permeability and porosity observed when subjecting cores to the simple processes used in these tests. Such are: hydration and swelling of clay, adsorption of ions, changes in surface structure and wettability, expansion and compression effects due to ice formation, shrinking and cracking, leaching of salts and colloids, displacement of particles resulting in either blocking or enlarging of pore openings. Whatever particular mechanisms are involved. however, it is apparent not only from this study but also from other investigations in the literature' not directly concerned with quick-freezing, that the effects produced by commonly used extraction, saturation and drying techniques may be of considerable magnitude The results of this study indicate that for the particular samples and techniques used, such effects are of the order of five to six times the effect of quick-freezing. insofar as changes in permeability are concerned. It may be argued that these samples might not include extremely shaly material where the effect of freezing upon permeability may be much greater. However, had such material been available for these tests, it would undoubtedly have been very susceptible also to alteration by the extraction and saturation treatment used. To investigate this point further, the individual sample data were re-grouped according to the magnitude of the average per cent permeability increases for the pairs of samples, irrespective of permeability. The results

Jan 1, 1953