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By Robert L. Greyer, John I. Myung

The Birdwell 3-D Velocity Log displays photographically the amplitude of elastic waves as a function of time and of the depth of the logging tool in a borehole. The elastic waves are generated by a magneto-strictive source transducer and detected by a receiving transducer, which is at a fixed distance from the source. The output of the receiver, which is proportional to the amplitude of the received elastic waves, modulates the light that exposes the film record. The duration of the receiving time can be selected to record the entire wavetrain of late events or to record only early events at a finer time scale. When a full complement of elastic waves (i.e., the refracted compressional wave, composite shear wave, direct fluid wave, and boundary wave) is recorded, their velocities and the density of the borehole fluid provide the information necessary to compute the elastic moduli and density of the rock surrounding the bore-hole. The differences between the values of the moduli determined from measurements of 3-D Velocity Logs and by other better known techniques are small, well within the limits of experimental error, when the velocity log is interpreted by an experienced seismic analyst. INTRODUCTION Quantitative descriptions of the mechanical properties of the rock environment have long been basic requirements of planning, design, and quality control of engineering projects. Because of this need, many techniques for determining the elastic moduli from static and

Jan 1, 1971

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

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

Jan 1, 1928

By Alan Stanley

THE MacIntyre Development of National Lead Co. is located at Tahawus, N. Y., in the heart of the Adirondack Mountains. Operations involve the mining and concentrating of a titaniferous iron ore to produce ilmenite and magnetite concentrates. A general description of the operation and metallurgy has been given by Frank R. Milliken.' Pigment plant production demands that the MacIntyre mill produce a 44.7 pct TiO, ilmenite concentrate. To achieve the required ilmenite grade and tonnage it is important that the table concentrate grade be closely controlled. Unfortunately, however, the titaniferous orebody which feeds the MacIntyre mill is not uniform. Ore dressing characteristics vary from one end of the orebody to the other, and from • one level to the next. The changeable nature of the mill feed precludes a single adjustment of the equipment for long periods of time. Thus the operators must constantly watch the equipment to insure a uniform concentrate from the fine and coarse tables and Wetherills, or dry magnetic separators. Chemical assaying of mill products requires about 4 hr from the time the sample is taken until assay results are obtained, and this is available only on a two-shift basis. The ore may change rapidly, even several times during a shift, so that assay results lose most of their control value by the time they are reported to the mill operating crew. Members of the crew have therefore tried to evaluate the table and Wetherill concentrate by visual inspection, since through long experience the shift operators, under most circumstances, can gage closely the grade of the mill products. However, there are times when the physical nature of the ore is radically different from normal. Under these conditions visual inspection is of no value, and at such times final ilmenite as low as 43 pct TiO, has been produced and shipped before the assay results have been received. The specific gravity method of assaying for TiO, has been attempted to eliminate the shipping of ilmenite below normal grade as well as to help control day to day and hour to hour mill production. Table I shows the minerals found in the MacIntyre ore along with their average weight proportions and specific gravities. The first two products considered for the specific gravity method were fine and coarse table concentrates. It was reasoned that these products were essentially ilmenite with the higher specific gravity gangue minerals. Since they were always produced the same way, and the desired grade of TiO, was always constant, the specific gravity of these materials would increase or decrease as the amount of ilmenite increased or decreased. Thus for table concentrates which assayed 40 pct TiOz a constant gravity would invariably be obtained, and as the TiO, value changed the specific gravity would change in direct proportion. The third product considered was Wetherill ilmenite. It was assumed that a desired grade of 44.7 pct Ti02 would also always contain the same amount and type of gangue minerals along with the ilmenite, and thus would always have the same specific gravity. As the TiO, value of the ilmenite concentrate changed so would its specific gravity. Dr. Kenneth Vincent, chief metallurgist of the Baroid Division of National Lead Co. at Magnet Grove, Ark., ran specific gravity tests on 17 samples of the desired products. The lowest specific gravity reading assayed the lowest in TiO, and as the specific gravity increased the trend was for the TiO, assay to increase, see Fig. 1. Since these results warranted further investigation, a 500-g capacity Torsion balance and 250 ml Le Chatelier specific gravity bottles were obtained. Shift samples of fine table concentrate, coarse table concentrate, and final ilmenite were tested. Each sample was split and 85 g weighed on the Torsion balance. The Le Chatelier bottle was filled with water to a zero mark. To avoid wetting the neck of the bottle it was found necessary to do this

Jan 1, 1953

By Alan Stanley

THE MacIntyre Development of National Lead Co. is located at Tahawus, N. Y., in the heart of the Adirondack Mountains. Operations involve the mining and concentrating of a titaniferous iron ore to produce ilmenite and magnetite concentrates. A general description of the operation and metallurgy has been given by Frank R. Milliken.' Pigment plant production demands that the MacIntyre mill produce a 44.7 pct TiO, ilmenite concentrate. To achieve the required ilmenite grade and tonnage it is important that the table concentrate grade be closely controlled. Unfortunately, however, the titaniferous orebody which feeds the MacIntyre mill is not uniform. Ore dressing characteristics vary from one end of the orebody to the other, and from • one level to the next. The changeable nature of the mill feed precludes a single adjustment of the equipment for long periods of time. Thus the operators must constantly watch the equipment to insure a uniform concentrate from the fine and coarse tables and Wetherills, or dry magnetic separators. Chemical assaying of mill products requires about 4 hr from the time the sample is taken until assay results are obtained, and this is available only on a two-shift basis. The ore may change rapidly, even several times during a shift, so that assay results lose most of their control value by the time they are reported to the mill operating crew. Members of the crew have therefore tried to evaluate the table and Wetherill concentrate by visual inspection, since through long experience the shift operators, under most circumstances, can gage closely the grade of the mill products. However, there are times when the physical nature of the ore is radically different from normal. Under these conditions visual inspection is of no value, and at such times final ilmenite as low as 43 pct TiO, has been produced and shipped before the assay results have been received. The specific gravity method of assaying for TiO, has been attempted to eliminate the shipping of ilmenite below normal grade as well as to help control day to day and hour to hour mill production. Table I shows the minerals found in the MacIntyre ore along with their average weight proportions and specific gravities. The first two products considered for the specific gravity method were fine and coarse table concentrates. It was reasoned that these products were essentially ilmenite with the higher specific gravity gangue minerals. Since they were always produced the same way, and the desired grade of TiO, was always constant, the specific gravity of these materials would increase or decrease as the amount of ilmenite increased or decreased. Thus for table concentrates which assayed 40 pct TiOz a constant gravity would invariably be obtained, and as the TiO, value changed the specific gravity would change in direct proportion. The third product considered was Wetherill ilmenite. It was assumed that a desired grade of 44.7 pct Ti02 would also always contain the same amount and type of gangue minerals along with the ilmenite, and thus would always have the same specific gravity. As the TiO, value of the ilmenite concentrate changed so would its specific gravity. Dr. Kenneth Vincent, chief metallurgist of the Baroid Division of National Lead Co. at Magnet Grove, Ark., ran specific gravity tests on 17 samples of the desired products. The lowest specific gravity reading assayed the lowest in TiO, and as the specific gravity increased the trend was for the TiO, assay to increase, see Fig. 1. Since these results warranted further investigation, a 500-g capacity Torsion balance and 250 ml Le Chatelier specific gravity bottles were obtained. Shift samples of fine table concentrate, coarse table concentrate, and final ilmenite were tested. Each sample was split and 85 g weighed on the Torsion balance. The Le Chatelier bottle was filled with water to a zero mark. To avoid wetting the neck of the bottle it was found necessary to do this

Jan 1, 1953

By AIME AIME

[The American Institute of -Mining Engineers does not assume responsibility for any statement of fact or opinion advanced in its papers or discussions.] A detailed list of the' papers contained in this section is given in the Table of Contents, pages i. and ii. Comments or criticisms upon all papers given in this section whether private corrections of typographical or other errors or communications for publication as "Discussions," or independent papers on the same or a related subject, are earnestly invited. ERRATA. The following corrections and. additions to papers contained in Bi-Monthly Bulletin No. 7, January, 1906,-were received too late for incorporation in the text of said papers ; but will be introduced in their: proper places: in the volume of the Transactions: These corrections are here published in order to bring them -promptly to the notice of the Institute. Page Line Cover, title-page and at the top of pages in the preliminary section "No. 1 " should be "No. 7 ". 1. Bottom For :" Dr. P. H. Shimer" read "Dr. P. W. Shinier " 69. Fig. 3. "N" of compass-record should be ?S". A new block for this illustration has been made as shown on p. xxxviii of this Bulletin. 78. 3 from bottom. For " slope " read " stope." 89. Fig. 2, small diagram in upper left-hand corner. Omit use 50 ft. to 1" scale ". For "Thro' 250 " read " Thro' 200 ". Add dotted line extend¬ing from upper right-hand corner of the square

Mar 1, 1906

By S. F. Emmons

The first paper which appears in the published Transactions of our Institute is that read by our respected Secretary at its first meeting in Wilkes-Barre in May, 1871. It is entitled " The Geographical Distribution of Mining Districts in the United States," and presents a brief but masterly review of what was known of the distribution of our deposits of useful minerals, particularly the metals, not only from a geographical but from a geological stand-point. At the request of Dr. Raymond I agreed, somewhat hastily, perhaps, to write for this occasion a brief sketch of the geological distribution of the deposits of the useful metals in this country, in the light of the increased knowledge of the present day. In the time given no personal investigation was possible, and as it was therefore out of the question to attempt to make anything that could be considered an original contribution to the history of our ore-deposits, I have been obliged to limit myself to an examination of such pub-

Jan 1, 1894

By R. T. Mishler

I WISH to express my keen appreciation of the article on the above subject by Coe and Clevenger.1 It has been doubly interesting to me, for the reason that the experience recorded and the principles evolved practically parallel the results of similar investigations made by myself and others at The Lucky Tiger Mine, El Tigre, Sonora, Mex. An outline of the work at El Tigre may aid in confirming the principles expressed by Coe and Clevenger. El Tigre has the unenviable distinction of possessing a slime with a much lower settling rate than any recorded in the article discussed. Under most favorable conditions the settling rate is about one-third that reported for the Liberty Bell mill, and one-tenth that reported for the Nipissing, Golden Cycle, and Presidio mills. This slow settling characteristic of Tigre ore has necessitated careful investigation of settling phenomena throughout the various evolutions of the plant. PREVIOUS SETTLING INVESTIGATION AT TIGRE During construction work at the mill, the Metallurgical Engineer in charge, D. L. H. Forbes, ran a series of experiments on settling. His tests proved that deep settling tanks were not ordinarily necessary. The use of tanks 2 or 3 ft. in depth was suggested.2 Later experiments demonstrated that the following principles of settling applied to Tigre pulp.3 1. In settling dilute pulp, settling rates are independent of pulp depth. 2. In settling thick pulp, settling rates are approximately proportional to pulp depth.4

Jan 3, 1917

By L. C. Brown, S. K. Behera

The diffusion rate in Ag-Sb couples is sensitive to con~pressive load with the width of Ag3Sb, the only phase present in the diffusion zone, increasing with stress up to 800 psi and remaining constant above this. Kirkendall marker experiments show silver to diffuse much faster than antimony in Ag3Sb and incipient porosity may therefore develop at the Ag/Ag3Sb interfnce restricting the transfer of atoms from the silver into the diflusion zone. Application of compressive stress reduces the tendency for porosity to develop and so increases the growth rate. In a recent paper Brown et al.1 observed a significant increase in the thickness of Cu2Te in Cu-Te diffusion couples on application of a compressive stress as low as 20 psi. Similar stress effects have also been observed in the Fe-A1,2 Al-u,3 arid cu-sb4,5 systems. It has been suggested that the increase in growth rates of intermetallic phases in these systems is due to a decrease in the amount of Kirkendall porosity with applied stress. In the present paper, results are presented of the effect of compressive stress on diffusion in Ag-Sb, together with a detailed examination of the Kirkendall effect. The Ag-Sb phase diagram6 shows that antimony has a moderate degree of solid solubility in silver, 5.7 at. pct at 350°C, but that there is essentially no solubility of silver in antimony. There are two intermediate phases— (hcp7) from 8.8 to 15.7 at. pct Sb and Ag3Sb (orthorhombic8) from 21.8 to 25.9 at. pct Sb. EXPERIMENTAL Diffusion couples were prepared from fine silver of 99.95 pct purity and from antimony of 99.7 pct purity. Both the silver and antimony were produced in the form of discs 1/2 in. in diam by approximately $ in. thick, with surfaces ground flat to 3/0 emery paper. Diffusion anneals were carried out in the apparatus previously described.1 A compressive load was applied to the diffusion couple through a lever arm system, with a reproducibility estimated to be ±10 psi. All runs were carried out in a protective hydrogen atmosphere. Following the diffusion anneal specimens were sectioned and polished and the width of the diffusion zone was measured metallographically. Composition profiles were measured using an electrostatically focused electron probe with a spot size of 10 , counting on Sb L radiation. Corrections for matrix absorptiori and fluorescent enhancement9 were not required. S. K. BEHERA, formerly Graducate Student, Department of Metallurgy, University of British Columbia, is now Postdoctoral Fellow, Whiteshell Nuclear Laboratories, Atomic Energy of Canada Ltd., Pinawa, Manitoba. L. C. BROWN, Junior Member AIME, is Associate Professor, Department of Metallurgy, University of British Columbia, Vancouver, B.C., Canada. Manuscript submitted June 14, 1968. IMD RESULTS Fig. 1 shows an electron probe traverse of a typical diffusion zone. In all couples examined only one intermediate phase was observed and the composition of this phase, 23 wt pct Sb, was in good agreement with the composition of Ag3Sb, 23 to 28 wt pct Sb. The presence of this phase was confirmed by X-ray diffraction of filings taken from the diffusion zone. The probe traverses showed no detectable solid solubility in either the silver or the antimony although the phase diagram indicates that some antimony, up to 6.5 wt pct pct Sb, should be in solid solution in the silver. However the width of this portion of the diffusion zone would be expected to be very small in view of the low diffusion coefficient in the silver, 4 x 10-l6 sq cm per sec at 350°C, 10 compared with that in the Ag,Sb, estimated as 3 x 10-8 sq cm per sec in the present work, and this region would therefore not be expected to be seen in the probe traverse. Application of stress resulted in a significant increase in the width of the diffusion zone, Fig. 2. At 350°C, the thickness of Ag3Sb increased from 250 at 0 psi to 400 p at the limiting stress of 800 psi, indicating an apparent 150 pct increase in the diffusion coefficient. Similar behavior was also observed at 400°C, indicating that the stress effect is not characteristic of just one temperature. The growth of Ag3Sb at 350°C and at various stresses is shown in Fig. 3. In every case the growth rate was parabolic indicating diffusion control. The kinetic curves all passed through the origin showing that delayed nucleation of Ag3Sb was not responsible for the stress effect and that it was a real growth effect. A series of tests were carried out in which diffusion was allowed to take place at a lower stress following an initial high stress diffusion anneal. Speci-

Jan 1, 1969

By G. O. Hiers

IN 1948 in the United States, 184,300 tons of lead was fabricated as coverings for electric power and communication cables. Such covering generally is called "sheathing" for the principal lengths of the cables, while the short lead encasements of the joints are called "sleeves." For the purpose of this paper, both lead sheathing and lead sleeving can fittingly be called lead tubing. In the United States in the same year, 34,262 tons of lead was consumed in making extruded lead pipes. Such pipe is employed frequently for water or chemical solution conveyance with gage pressures up to 100 psi. Consequently, the walls are thicker in ratio to diameter than is the case with the cable sheaths. For this discussion, lead pipes can be thought of as tubes or tubing. About 19 centuries ago, the Romans produced metallic lead of about 99.9 pct purity. Evidently this was cast in flat sheets of Y8-in. thickness, for instance. Such sheet of 10-ft lengths was bent around a removable core producing a tube somewhat oval in cross-section and with a typical internal diameter of, say, 3 in. This method of production necessitated a seam or longitudinal joint evidently made by casting the seam metal adherently. The seam metal, as analyzed recently, contains about 1 ½ pct Sn. Blow pipes were said to have been employed when the joint was made. Specimens of such pipe were found both in Italy and in the British Isles. It is believed the pipes conveyed water for drinking purposes and for baths. Numerous specimens are still in existence in a good state of preservation, due to lead's usual inertness in air or water. According to Claude E. Pearson,1 there were various stages in the evolution of lead extrusion presses for making tubes: I. First press by Joseph Bramah, 1797. 2-First successful press by Thomas Burr, 1820. 3-First bridge die press, J. and C. Hanson, 1837. In 1845, Wheatston and Crooke made the first lead sheathed electric cable. Flat lead strip was wound spirally around the insulated conductor. Overlapping edges of lead were soldered. Early cables were also made by pulling insulated conductors into 50-ft lengths of lead tube. Borel in France, and Wesselau are credited with devising the first presses for extruding lead sheaths directly on cables in 1879. In the United States, Eaton, in 188o, and Roberston, in 1885, introduced new presses of improved design. Following Cunningham's original idea of making lead traps and. bends in 1873, the presses apparently were improved and made successfully soon thereafter. Between the years 1930 and 1940, two continuous extrusion presses for lead were developed abroad. One apparently was originated by Dunsheath and made by the Henley Telegraph Works Co. The other was developed by the Pirelli-General Cable Co. CURRENT PRODUCTION In addition to the several descriptive chapters on lead extrusion in Pearson's book, the Metals Handbook2 recently described pipe, sheath, and collapsible tube production. The latter is by impact extrusion.

Jan 1, 1951

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On' (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By H. W. Straley

Some ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbeet.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information On' (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college "algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1946

By J. P. Manning

UNQUESTIONABLY, the outstanding feature of the piping for the sulphide plant is the large amount which had to be done in almost every size from instrument tubing to 84 in. OD pipe. In this article the piping will be considered under the main headings of Yard Piping and Building Piping. Tables I and II give data concerning the individual systems. In the case of yard piping the figures pertaining to lengths and weights are approximately correct; those for building piping are given only for the concentrator and smelter buildings where the amounts were the most extensive. Yard Piping The installation specifications for yard piping called for all sizes larger than 12 in. to have welded joints made with outside welding rings. This type of joint was selected over butt-welded joints because less time would be needed for lining up, damaged pipe ends would be less of a problem, and deflections up to 6° per joint could be made when necessary. These advantages were thought to outweigh the disadvantage of the greater amount of welding required. Yard piping in sizes of 3 to 12 in. was specified to have butt-welded joints made with inside welding rings which were intended to facilitate lining up and keep welding material out of the lines. Screwed connections were specified for sizes under 3 in. No protective coatings or wrappings were required on pipe to be buried because of the dry climate at Chuquicamata.

Jan 1, 1952

By R. S. Silo

A new process for leaching oxidized copper ore (CuO) is presented. Its principle advantage is that it requires low cost raw materials. For large operations the only materials used are sodium chloride (NaCl) and calcium carbonate (CaCO3). In those operations of medium size, sulfur, which is abundant in the North of Chile, is used. However the process may also be applied to small copper plants provided low cost sulfuric acid is available as a substitute for sulfur. That is, it may be applied to those small capacity plants where large additional installations for the regeneration of the leaching solution are not needed. The method is comprised of six stages, namely: Crushing, Leaching, Washing of the Solution, Precipitation, Regeneration of the Leaching Solution, and Fusion. A flow diagram for this process in a large or medium-sized plant is given. LEACHING This process requires hydrochloric acid as the leaching solution for treating copper ores. This is easily produced from sulfuric acid and sodium chloride as follows: A small excess of NaCl is advisable to avoid the presence of free sulfuric acid. The experiments reported here were performed with 36 grams of HC1 per liter of solution to obtain a free acidity equivalent to 50 gpl of H2SO4 (the usual leaching solution concentration). Hydrochloric acid leaching may be done by either of the usual methods; i.e. percolation or agitation. The method selected depends on the characteristics of the ore. No matter which method is chosen, low pressure air can be used either as an air lift for circulation of the solution, or for agitation of the pulp with "Pachuca" tanks. In either case, there are no problems with respect to oxidation. Leaching by agitation produces good recovery; but at times, the problem of high consumption of leaching solution or the overproduction of grinding slimes which interfere with contact between phases negates this advantage. The percolation method was used in this study. Percolation took place in four 15-liter containers, 11 1/2 in. diam. by 12 in. height, having hemispheric bottoms. Each container was fitted with a wooden bottom having 1/4 in- holes. The air lift was a 3/4 in. plastic tube 10 in. high. Air at 6 psi was supplied from a rotating laboratory compressor. 10 kg of ore were charged to the container and counter-current leaching was used. The leaching solutions were fed through the center pipe in order to use the ascending solutions to reject the air in the charge. The solution circulated through the air lift in order to facilitate the movement according to the increase in density produced by the reaction between the phases present. Specific observations made in this report cannot be generalized because the leaching characteristics are dependent on the nature of the ore used, the method of leaching, and the production of primary or secondary slimes while crushing or leaching.

Jan 1, 1967

By H. Conrad, V. V. Damiano, G. J. London

A method is described for producing single crystals of high-purity beryllium, Be-4.37pct Cu, and Be-5.24 pct Ni. These crystals were prepared for testing in compression parallel to the [0001] by orienting and lapping to within ±3' of arc of the (0001). Microstrain testing apparatus is described along with c axis compression results for ingot purity beryllium, twelve-zone-pass material, and the above-mentioned alloys. Results show no measurable plasticity for the ingot purity material from -196" to 400°C, although some surface traces of (1122) slip was observed at 200°C and above. The twelve-zone-pass material shows substantial microstrain plasticity at 220°C with slip on (1122). Both alloys show significant plasticity at room temperature and above with slip also on (1122) planes. THE two slip systems which normally operate during the plastic deformation of beryllium in the vicinity of room temperature are:' basal slip (0001)(1120) and prism slip . Pyramidal slip with a vector inclined to the basal plane has been reported for elevated temperatures,'-a but occurs near room temperature only at very high stresses.~ A summary of the available data on the effect of temperature on the critical resolved shear stress for slip on these systems has been compiled by Conrad and Perlmutter.~ It has been postulated6'7 that one of the principal factors contributing to the brittleness of poly crystalline beryllium at temperatures below about 200°C is the difficulty of operating pyramidal slip with a vector inclined to the basal plane. Hence, detailed information on the operation of such a slip system is important to understanding the brittleness of beryllium. The operation of pyramidal slip with a vector inclined to the basal plane is best accomplished in beryllium by compressing single crystals in a direction parallel to the c axis. In such a test the resolved macroscopic shear strzss on the basal and prism planes is zero and (1012) twinning which is favored by tension along the c axis does not occur. Hence, in c axis compression of beryllium the normal deformation modes are inhibited and the operation of pyramidal slip with a vector inclined to the basal plane is favored. In the present investigation, c axis compression tests were performed on beryllium single crystal as a function of temperature (77" to 700°K), purity (commercial and twelve zone pass), and alloy content (4.37 wt pct Cu and 5.24 wt pct Ni). Presented here is a description of the test techniques employed and the gross mechanical behavior observed. A detailed analysis of the slip traces developed on the surfaces of the deformed specimens during these tests and the results of electron transmission studies of the deformed crystals are given in a separate paper.B PROCEDURE 1) Materials and Preparation. Single crystals about 1 in. diam were prepared of the following materials: commercial-purity beryllium, high-purity beryllium, and two beryllium alloys, one with 4.37 wt pct Cu and the other with 5.24 wt pct Ni. The commercial-purity single crystals were obtained by cutting specimens from large-grained ingot of Pechiney SR material, which is approximately 99.98 pct pure. The high-purity crystals were prepared by floating-zone refining (twelve passes) a rod (7 in. by 1 in, diam) of Pechiney SR grade cast and extruded beryllium. Although an absolute chemical analysis of the zone-refined material was not established, mass spectro-graphic analysis, emission spectrographic analysis, and y activation analysis indicated that it contained in atomic fractions about 5 to 10 ppm each of carbon and oxygen, 1 to 5 ppm each of nickel and iron, and about 1 to 2 ppm of copper, with the remaining residual impurities being less than 1 ppm. Further indication of the purity of this material is provided by the critical resolved shear stress for basal slip, which was approximately 300 psi. The starting material for the alloy single crystals was 1-in.-diam floating-zone-refined (six passes) rod of Pechiney SR grade beryllium. Two such rods were wrapped respectively with sufficient weight of wire of high-purity copper (99.999 pct) or nickel (99.999 pct) to yield a 5 wt pct alloy. A seventh floating-zone pass was then applied to each of the rods to accomplish the initial alloying and an eighth pass for homogenization. Analytical samples were taken from regions of the rod immediately adjacent to where the mechanical test specimens were cut; these indicated 4.37 wt pct Cu and 5.24 wt pct Ni. 2) Crystal Orientation. To avoid the occurrence of basal slip during c axis compression testing, it is necessary to load the crystals as nearly parallel to the c axis as possible. Preliminary c axis compression tests indicated that plastic flow and/or fracture occurred at stresses of the order of 300,000 psi; hence on the basis of a critical resolved shear stress for basal slip of 300 to 400 psi, the maximum crystal misorientation permitted is about 4 to 5' of arc. Since this accuracy cannot be obtained using the usual back-

Jan 1, 1969

By B. W. Frazier

IN the numbers of the Engineering and Mining Journal of June 27th and July 11th, 1874, there appeared some very complete statistics of the working of some anthracite blast-furnaces belonging to a large establishment in this country, followed by calculations of the heat requirements of the furnaces, similar to those which have been made for European furnaces by Mr. I. Lowthian Bell, Prof. Gruner, Prof. Akerman and others. To the author of these valuable papers, Mr. John A. Church, of New York, the thanks of American iron metallurgists are due for this attempt (the first, to my knowledge at least), to attack systematically and scientifically the problem of economy of fuel in our anthracite blast-furnaces, by applying the formulas obtained from European practice and from the experiments of European metallurgists, to the determination of the amounts and the modes of distribution of the heat production and absorption in our furnaces. In order to facilitate a comparison of the results thus obtained for the anthracite furnaces in question with those obtained for the best types of European coke furnaces, I have reduced to the same standard of heat units (the product of the ton and the degree Centigrade), and placed side by side, the calculations of the heat requirements of one of the furnaces of the Clarence works, as given by Mr. Bell, of a furnace at Pouzin, on the banks of the Rhone, given by Prof. Gruner, and of the anthracite furnaces for 1869-73, as given -by Mr. Church. Table I shows the different conditions of working of the three furnaces, of which Table II gives the respective heat requirements. I have made this latter table in such a manner as to give for each furnace its heat requirements, both according to the views of Mr. Bell and those of Prof. Gruner. Mr. Bell assumes that all the carbonic acid produced by the decomposition of the carbonate of lime of the charge, is reduced to carbonic oxide by contact with solid carbon ; also, that the carbon of the pig iron is a portion of that deposited in the ore in the upper part of the furnace, by the dissociation of carbonic oxide. Prof. Gruner, on the other hand, assumes that all the carbonic acid of the carbonates of the charge escapes unreduced, and that the carbon deposited in the upper part of the furnace, by the dissocia¬tion of carbonic oxide, is again converted into carbonic oxide or

Jan 1, 1875

By O. C. Wheeler

The reports for the years 1942 and 1943 were written in 1943 and 1944, respectively, and were summarized at the meetings of the Institute in February of those two years, but the manuscripts were held until released for publication in 1945. 1942 Lack of transportation facilities during 1942 resulted in a sharp curtailment of producing operations in Colombia at the middle of the year, and production of crude for export was thus greatly affected. Total production in 1942 amounted to only 10,617,008 bbl. Of this amount the Tropical Oil Co. produced 9,453,995 bbl., which includes 126,607 bbl. of petroleum condensate, and the Colombian Petroleum Co. produced 1,163,013 bbl, from its Barco property. The following summary, showing a comparison of the number of barrels of crude produced during the years 1941 . and 1942, reflects the extent to which producing operations were reduced during 1942: These above figures do not include any production recovered from Condor's (Shell's) Casabe wells during testing operations. An active exploratory drilling Program was carried out by several of the major companies, and the year 1942 established a record for wildcatting activity in Colombia. Excluding development operations on the Shell's Casabe wells on the Yondo concession, 13 wildcats were completed. Eight of these were in the Magdalena Valley, and all were abandoned; five were on the Barco Concession, and of these four were completed as producers. Particularly notable among the petroleum developments during 1942 was the continued success of drilling operations on the Shell's Yondo Concession of Compania Colombiana de Petroleos El Condor opposite Barranca-Bermeja. In 1941 the discovery well, Casabe No. I, was drilled and completed with an estimated initial production of approximately 480 bbl. of 19 1/2 A.P.I. gr. oil per day. In 1942, seven stepouts from the discovery well were successfully completed with initial production ratings ranging from 150 to 1230 bbl. per day. An extensive wildcat program on the Barco Concession of the Colombian Petroleum Co. yielded favorable results with the completion of four producers in the Socuavo and Tres Bocas areas; a fifth test, Sardinata Norte No. I, was abandoned at 500;" feet. Wildcatting operations in the lower Magdalena Valley area were initiated in 1942 With the drilling of a deep test, El Dote No. I, by Compania de petroleo Shell de Colombia. The well was a failure and was abandoned at a depth of 10,046 (?) feet. Wildcats on which drilling operations were being conducted at the end of the

Jan 1, 1945

By W. M. Courtis

In compliance with the suggestion in Mr. Emmons's paper on the Genesis of Certain Ore-Deposits (Trans. xv., 125), that facts should be accumulated for study and future compilation, I wish to put on record, before the data are lost, a section of the Animikie series of rocks and the vein-phenomena exposed by the explorations at the Duncan mine, Port Arthur, Thunder Bay, Lake Superior. This may serve as a supplement to my paper on this district (Trans. v., 473). I trust this information will not be without interest at. the present time. The recent discoveries of very rich silver-ores at Rabbit and Silver Mountains, on the same belt, have brought the district once more before the public; and a study of the facts here set forth may assist some fellow-engineer in the determination of the value of property offered for sale in this neighborhood. The vein at the Duncan mine was tested to a depth of over eight hundred feet by shaft and drifts, and to nearly a thousand feet with

Jan 1, 1887

By H. W. III Straley

SOME ten years ago the writer1 made a survey of college catalogues to determine what sort of training geologists were receiving in basic sciences. In the light of this compilation and subsequent experience, he made the statement at the 1939 meeting of the Institute that apparently only in schools offering a curriculum in geological engineering or in the geological option of mining engineering curricula could geologists hope to receive the training advocated by Hubbert.2 In the former study, schools of engineering were excluded because the preparation was known beforehand. The present study aims to consider engineering schools in an effort to determine not only what basic studies are required but what use is made of them in geological courses of the third and fourth years. The present study combined the method of consulting catalogues with a questionnaire. Letters were mailed to officials responsible for supervision of mineral industries curricula requesting information on: (I) the extent to which the engineering preparation of the first two years is actually used in the work of the last two; (2) the extent to which the engineering of the last two years is used in courses in geological engineering or geology; and (3) whether geologists or geological engineers teach descriptively or analytically. Both methods of securing information are open to criticism; the former on the basis of the incomplete information supplied by university catalogues concerning prerequisites and curricula and the latter on the ground that the questions may be phrased poorly or given insufficient consideration by replying officials. Fifty-six questionnaires were mailed and 39 (70 per cent) replies were received. Of the replies, 6 (11.3 per cent) were from colleges of liberal arts unconnected with engineering schools, 11 (20 per cent) were from schools having curricula in geological engineering and 22 (39 per cent) were from departments of mining engineering. The number of replies and their distribution were satisfactory in spite of the receipt of only six from liberal arts colleges. It should be understood that the universities of which the catalogues were examined were not necessarily the same in both compilations. No departments that are admittedly or known to be acting solely as service departments were recorded in the second study. Table I shows that all professional schools require mathematics through calculus, whereas only 7 per cent of the liberal arts curricula expect the same level of preparation; only 73 per cent require trigonometry and 40 per cent college algebra. The proportion of geological engineering curricula requiring differential equations is higher than mining. All of the engineering schools require general and analytical chemistry, whereas only 77 per cent of liberal arts curricula require general and 18 per cent analytical. Twenty-two per cent of the mining and 40 per cent of the geological engineering

Jan 1, 1941

By C. E. McCarthy, A. George Stern, Stanley J. Green

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948

By A. George Stern, Stanley J. Green, C. E. McCarthy

The national interest prompts consideration of any new source of mineral wealth even though the immediate need may be of minor importance. A critical shortage of potash in the United States during the World War of 1914-18 stimulated broad development, which established an American potash industry on a basis so sound that this commodity is not even listed as urgent in the present war program. However, this should not minimize its importance, and the need remains, not only for adequate supplies of the usual potash salts (the chloride and the sulphate) but of all potash chemicals. Potassium carbonate (known commercially as pearl ash) has been a high-cost commodity; in consequence, the quantity used by industry has been relatively small. By a combination of fortuitous geographical circumstances, there is opportunity to produce a low-cost potassium carbonate with the possibility that this might become an important source of chemical potash. The following paper considers a recently developed process1 taking advantage of these conditions. Wyomingite Deposit For many years it has been known that a deposit of potash-bearing igneous rock known as wyomingite occurs in the south- western part of Wyoming at Zirkel Mesa, which is 44 miles northeast of the town of Green River (Fig. I). The western escarpment of this deposit is about 3 miles east of the Superior Branch of the Union Pacific Railroad. This mesa alone covers approximately 3300 acres, and the average thickness of the wyomingite flow is 75 ft. It has been estimated2 that about one billion tons of wyomingite occur in this one deposit, all of which can be mined by open quarrying without overburden. In addition, there are several smaller deposits within 10 miles, so that the aggregate wyomingite in the area approaches two billion tons. This is one of the world's largest sources of leucite, and the only known deposit of its kind in the United States. The wyomingite (which is in the form of a lava flow and numerous small craters), with the mineral leucite as its most important component, has a mineral composition3 of: leucite, 50 per cent; phlogopite, 15; diopside, 12; kataphorite, 15; glass and apatite, 8. The phlogopite is present in visible phenocrysts 1/8 to 1/2 in. in diameter, but the leucite usually is almost microcrystalline and not capable of liberation by grinding. Chemically, the Wyomingite has a potash Content averaging slightly higher than eleven per cent K2O. The potash is mainly in the leucite, but part of it is in the phlogopite and a small amount is in the feldspars that are occasionally seen as undigested occlusions of the granite country rock, through which the leucite-bearing lava intruded, and as sanidine feldspar both in the crystalline and glassy phases of the rock. A typical analysis of the wyomingite from Zirkel Mesa is as follows:

Jan 1, 1948