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By M. A. Mahtab, T. M. Yegulalp

This paper presents the development and application of an approach for clustering fracture orientation data. Data are projected on the surface of the unit upper hemisphere and the clustering approach is applied in three steps: (1) randomness test and identification of dense clusters, (2) estimation of cluster statistics, using Bingham distribution, and (3) rejection of extremal data based on Chauvenet's criterion.

Jan 1, 1982

By George S. Rice

INTERESTING developments going on in European coal mines look to: (1) increasing mechanization; 12) concentration of mining; (3) improvement in safety appliances; and (4) studies in bettering roof supports and. incident to this, studies of the way a roof behaves under different conditions of support and methods of mining and back-filling. Mechanization and concentration off mining is being adapted to a considerable extent in Great Britain coincident with the development of conveyors. Belt conveyors, shaker-pan conveyors, and a combination of the two are used. For pitching seams not steep enough for the coal to run in station-chutes, the. shaker-pan conveyor is more popular. For large outputs from a contiguous face, the double-unit system of face layout-that is, the conveyors meeting at a common point-is favored.

Jan 1, 1934

By W. N. Poundstone, P. C. Thakur

Introduction Horizontal drilling in coal mines is a relatively new technology. The earliest recorded drilling in the United States was done in 1958 at the Humphrey mine of Consolidation Coal Co. for degasification of coal seams. Spindler and Poundstone experimented with vertical and horizontal holes for several years. They concluded in 1960 that horizontal drilling in advance of underground mining appeared to offer the most promising prospect (for degasification) but effective and extensive application would be dependent upon the ability to drill long holes, possibly 300 to 600 m, with reasonably precise directional control and within practical cost limits (Spindler and Poundstone, 1960). Mining Research Division of Conoco Inc., the parent company of Consolidation Coal Co., began a research program in the early 1970s to achieve the above objective. The technology needed to drill nearly 300 m in advance of working faces was developed by 1975 and experiments on advance degasification with such deep holes began in 1976. Preliminary results of this research have already been published (Thakur and Davis, 1977). To date nearly 4.5 km of horizontal holes have been drilled for advance degasification and earlier results were reconfirmed. In summary, these are: • The greatest impact of these boreholes was felt in the face area where methane concentrations were reduced to nearly 0.3% in course of two to three months from original values of nearly 0.95%. • The methane concentration in the section return reduced to 50% of its original value immediately after the boreholes were completed, indicating a capture ratio of 50%. • The total methane emission in the section (rib and face emission plus the borehole production) did not increase but rather gradually declined with time. • Initial production from 300 m deep boreholes in the Pittsburgh seam varied from 3 m3/min to 6 m3/min but then slowly declined as workings advanced inby of the drill site (well head) exposing a larger surface area parallel to the borehole. Encouraged by these results, it was decided to design a horizontal drilling system that would be mobile and compatible with other face equipment. A mobile horizontal drill can be divided into three subsystems: the drill rig, the drill bit guidance system, and borehole surveying instruments. The drill rig provides the thrust and torque necessary to drill 75- to 100-mm diam holes up to 600 m deep and contains the mud circulation and gas cuttings separation systems. The drill bit guidance system guides the bit up, down, left, or right as desired. Borehole surveying instruments measure the pitch, roll, and azimuth of the borehole assembly. Additionally, it also indicates the thickness of coal between the borehole and the roof or floor of the coal seam. Thus, it becomes a powerful tool for locating the presence of faults, clay veins, sand channels, and the thickness of coal seam in advance of mining. In recent years, many other potential uses of horizontal boreholes have come to light, such as in situ gasification, longwall blasting, improved auger mining, and oil and gas production from shallow deposits. The purpose of this paper is to describe the hardware and procedure for drilling deep horizontal holes. The Drilling Rig [Figures 1 and 2] show the two components of the mobile drilling rig: the drill unit and the auxiliary unit. The equipment (except for the chassis) was designed by Conoco Inc. and fabricated by J. H. Fletcher and Co. of Huntington, WV. The drill unit. It is mounted on a four-wheel drive chassis driven by two Staffa hydraulic motors with chains. The tires are 369 X 457 mm in size and provide a ground clearance of 305 mm. The prime mover is a 30-kw explosion-proof electric motor which is used only for tramming. Once the unit is Crammed to the drill site, electric power is disconnected and hydraulic power from the auxiliary unit is turned on. Four floor jacks are used to level the machine and raise the drill head to the desired level. Two 5-t telescopic hydraulic props, one on each side, anchor the drill unit to the roof. The drill unit houses the feed carriage, the drilling console, 300 m of 3-m-long NQ, drill rods, and the electric cable reel for instruments. The feed carriage is mounted more or less centrally, has a feed of 3.3 m, and can swing laterally by ± 17°. It can also sump forward by 1.2 m. The drill head has a "through" chuck such that drill pipes can be fed from the side or back end. General specifications of the feed carriage are: [ ] The auxiliary unit. The chassis for the auxiliary unit is identical to the drill unit but the prime movers are two 30-kW explosion proof electric motors. It is equipped with a methane detector- activated switch so that power will be cut off at a preset methane concentration in the air. No anchoring props are needed for this unit. The auxiliary unit houses the hydraulic power pack, the water (mud) circulating pump, control boxes for electric motors, a trailing cable spool, and a steel tank which serves for water storage and closed-loop separation of drill cuttings and gas.

Jan 1, 1981

By Edward N. Trump

Extensive beds of rock salt occur in New York, Michigan, Kansas, and Texas. Wells are drilled through the beds, cased, and equipped with a suspended center tube. By circulating water through such a well, the soluble saline is dissolved and a brine produced. Operation of a well results ordinarily in the formation of an inverted conical cavity in the saline, inasmuch as the dissolving action of the fresh water is more rapid at the top of the bed than at the bottom. Under continued operations, the side wall of the cavity becomes so flat that a blanket of mud is deposited thereon and if the bed is overlain with shale or a soft rock caving follows. Three methods of operating wells have been employed—the "Tully plan," "the Detroit plan," and "the Trump plan." Both the Tully plan and the Detroit plan involve the formation of the aforesaid conical cavity with constant caving of shale above and deposition of mud on the side wall; the Trump plan was developed to avoid both these circumstances and thus materially prolong the life of a well. Details of the three plans follow. The Tully Plan In the Tully plan, the water passes down the annular space between the casing and the center tube, the brine rising in the center tube (Figs. I and 2). An inverted conical cavity is formed, the side wall of which relatively soon becomes so flat that it is blanketed by mud and impurities from the rock salt; this blanket limits dissolving action to the perimeter of the base of the cone, situated at the top of the saline bed. Caving of the overlying strata follows, and the total production of brine from the well is comparatively small. The Detroit Plan In the Detroit plan, the water passes down the center tube, the brine rising in the annular space between the center tube and casing (Figs. I and 2). Initially, this flow forms a pear-shaped cavity in the bottom of the salt bed, but this cavity soon changes in shape to that of an inverted cone, as under the Tully plan, with the attendant disadvantages inherent in that plan. Application of the Detroit plan is limited to salt beds 300 ft. thick as a minimum. This depth of bed is necessary to prevent fresh water of lower specific gravity than the brine from rising in the

Jan 1, 1948

By Edward N. Trump

Extensive beds of rock salt occur in New York, Michigan, Kansas, and Texas. Wells are drilled through the beds, cased, and equipped with a suspended center tube. By circulating water through such a well, the soluble saline is dissolved and a brine produced. Operation of a well results ordinarily in the formation of an inverted conical cavity in the saline, inasmuch as the dissolving action of the fresh water is more rapid at the top of the bed than at the bottom. Under continued operations, the side wall of the cavity becomes so flat that a blanket of mud is deposited thereon and if the bed is overlain with shale or a soft rock caving follows. Three methods of operating wells have been employed—the "Tully plan," "the Detroit plan," and "the Trump plan." Both the Tully plan and the Detroit plan involve the formation of the aforesaid conical cavity with constant caving of shale above and deposition of mud on the side wall; the Trump plan was developed to avoid both these circumstances and thus materially prolong the life of a well. Details of the three plans follow. The Tully Plan In the Tully plan, the water passes down the annular space between the casing and the center tube, the brine rising in the center tube (Figs. I and 2). An inverted conical cavity is formed, the side wall of which relatively soon becomes so flat that it is blanketed by mud and impurities from the rock salt; this blanket limits dissolving action to the perimeter of the base of the cone, situated at the top of the saline bed. Caving of the overlying strata follows, and the total production of brine from the well is comparatively small. The Detroit Plan In the Detroit plan, the water passes down the center tube, the brine rising in the annular space between the center tube and casing (Figs. I and 2). Initially, this flow forms a pear-shaped cavity in the bottom of the salt bed, but this cavity soon changes in shape to that of an inverted cone, as under the Tully plan, with the attendant disadvantages inherent in that plan. Application of the Detroit plan is limited to salt beds 300 ft. thick as a minimum. This depth of bed is necessary to prevent fresh water of lower specific gravity than the brine from rising in the

Jan 1, 1948

By F. N. Speller

ASIDE from the Tuesday symposium on gases in metals, the Iron and Steel Division swung into action for the first time on the third day of the meeting with a discussion of iron and steel alloys. Dr. E. S. Davenport opened the discussion on Wednesday morning by presenting the paper written by himself in association with E. C. Bain and W. S. N. wiring on the equilibrium diagram of iron-manganese-carbon alloys. A written discussion by Mr. Scott showed qualitative agree- ment, but some quantitative disagreement. Messrs. Marsh and Sisco discussed the paper at length, pointing out non- agreement of certain conclusions with the phase rule. Dr. Krivobok, although differing with the authors on some points, stated that in his opinion the information collected would be valuable in the practical use of these alloys.

Jan 1, 1932

By M. E. Wadsworth, R. L. Pearson

THE dissolution of UO, in carbonate solutions can occur only il the tetravalent uranium is oxidized to the hexavalent state. The carbonate system is of particular interest because it provides a means whereby the oxidized uranium may be dissolved by forming the complex ion UO,(CO3)3-1. In the presence of oxygen the over-all reaction may be written UO+1/2O,4CO3T+2HCO2UO,(CO4)11+H2O [1] UO, was used in this study because it could be obtained in high purity and also because it is representative of the most refractory of the primary uranium minerals. The mineral ulrichite is the natural counterpart of the synthetic UO, used. The kinetics of the leaching of UO, should be similar for any of the uraninite type minerals. A pressurized system was used not with the hope it could be extended to commercial systems as a primary purpose but simply to provide the means whereby important temperature and pressure parameters could be varied for the evaluation of the kinetic processes. The importance of such pres-

Jan 1, 1959

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 John Fritz

THE John Fritz Medal was presented to Ambrose Swasey of Cleveland, Ohio, in the Auditorium of the Engineering Societies Building, in New York, on April 23. Charles F. Rand, Chairman of the Board of Award, presided and, after speaking very briefly of Mr. Swasey, whose friend he has been for a great many years, ,introduced Gen. William Crozier, formerly Chief of Ordnance, U. S. Army, who spoke most entertainingly of, Mr. Swasey's accomplishments in designing and manufacturing instruments of precision, particularly for the use of artillery in range finding: Dr. William Wallace Campbell, Director of the Lick Observatory, and President of the University of California, described Mr. Swasey's, contribution to astronomical advancement by his notable development of the largest telescopes in this country.

Jan 1, 1924

By Scott Turner

AS IN OTHER lines of engineering, progress in mining was influenced during 1931 by the world-wide economic depression. Low-metal prices ? resulted in active efforts to reduce production costs of base-metal mines. Probably the chief factors in reducing costs were the greater efficiency of labor, due to better relations between management and workmen, to the scarcity of jobs, reduced wages in most base-metal mines, and to lower overhead expenses secured by drastic economies. The use of mechanical equipment continued to increase, due to more efficient designs and a better practical understanding of the applicability of such equipment to requirements. With one exception, there were no startling innovations in stoping methods, but there was notable progress in the better adaptation of mining methods. In the United States, no large new base-metal or silver mines came into production during the year. At a number of mines, output was curtailed, and operations were discontinued at several important producers. On the other hand world-wide interest in gold was stimulated, as evidenced by increased output from producing mines, the reopening of some old properties, and intensive prospecting for new deposits. Mining methods at several gold mines in the United States and Canada were modified or changed to obtain a higher extraction of ore or to meet the need for additional support of the mine workings at depth. Permanent record of improvements in mining practices has been made during the year by the Federal Government; the industry thus has available for its use more detailed information than ever before. This work was done through cooperation of the mining companies and their representatives with the United States Bureau of Mines, whereby many papers descriptive of practices and costs at individual mines have been published. The Bureau's program of describing methods and costs at the more important mines (those producing in excess of $100,000 annually) is nearly completed, but papers will be issued concerning such additional mines as may possess particular interest because of novel methods, exceptional conditions or operating results, and others to supplement reports already issued. In this way, an interesting record of the results of changing conditions, or innovations in methods, will be made available.

Jan 1, 1932

By G. R. Fitterer

Trends in slag composition in acid open hearth practice, particularly the variation in iron and manganese oxides during refining, are reviewed. A procedure which is currently being used to control the refining reactions is described and partially explained through slag phase diagram considerations. In addition, the ionic nature of these slags is considered and it is proposed that silica combines with the oxide ion and forms a series of complex silicate ions. The effect of basic oxides such as FeO, MnO, and CaO is apparently additive in furnishing the oxide ion for this purpose. IT is now possible to control192 the refining of steel in an acid open hearth furnace to a surprising degree. This control has been brought about through the knowledge and use of several facts and tests. These are: 1—A planned charge analysis based on a knowledge of the type of fuel, the bath depth, etc. 2—Knowledge and control of rate of fuel input as well as its method of atomization, if any. 3—Determination of the slag fluidity. 4—Bath temperatures (use of the Pt-Pt-Rh thermocouple). 5—Periodic determination of carbon content (use of the Carb-analyzer or Carbometer). Through the use of these control methods, a heat may be made to "melt-in" with or without residual manganese and silicon. The rate of refining or carbon elimination may be controlled accurately, and the desired tapping (or "go-ahead") carbon may be approached at a rapid or slow rate according to preference and type of steel being made. Thus operating time may be predetermined and greatly reduced. Further, all of the refining can be done without any ore or other oxidizing additions, and the heat may be brought to the desired tapping temperature simultaneously with the attainment of the desired carbon analysis. This control procedure is being used in most acid plants in the United States today and has been de- scribed.', ' Although the slag metal relations are sufficiently well known to control the heat as stated, yet it is possible to account for only about two-thirds' of the 0, which must have been used for the elimination of the carbon and other metalloids. Phase diagram studies' of the slags, slag weights, the metal particle theory,' ore additions, limestone additions, etc., may all be taken into consideration; yet, the mechanism of oxidation and source of all the oxygen remain to be explained. This paper reviews some of the principles upon which the control methods were based and proposes other possibilities regarding the unexplained phenomena. Old and New Acid Open Hearth Practice Some of the recent statements in the literature regarding acid open hearth operation are based on old heat data which are not typical of modern American practice. These old heat logs have been responsible for many misconceptions regarding the acid hearth practice and represent heats which do not coincide with modern methods. In recent years, American practice has been altered materially and these changes will be briefly described. Modern American Practice: A typical log of a modern American acid open hearth heat' is shown in Fig. 1 which indicates that the refining time was only two hours, during which the carbon was eliminated rapidly and the bath temperature increased at a fast rate. The manganese and silicon contents of the metal were essentially eliminated and their oxides had become constituent parts of the slag at the "melt-down" period. During the course of the heat, the MnO content of the slag remained practi-

Jan 1, 1954

By Thomas V. Falkie, Robert Stefanko

Recently The Wall Street Journal featured a series of articles titled "The Dirty Work-Brutal, Mindless Labor Remains a Daily Reality for Millions in The US.-Mining Coal, Shoveling Slag, Gutting Hogs Pays Bills But Can Drain Spirit, Hope." The article was devoted to describing how tough and dangerous is mining (and some other jobs) and how rough life is for miners and their families. Similarly during the past few years some psychiatrists and sociologists have devoted increased attention to the attitudes and life styles of miners. Unfortunately most of their approaches have been neither scientific nor objective, and very little good clinical research has been done on the subject.

Jan 1, 1971

By Charles W. Goodale

The principles and theory on which the U. S. mining law of 1872 was based are well understood, and have been discussed at great length by many writers. The papers by Dr. R. W. Raymond1 in the Transactions of the American Institute of Mining Engineers have covered the subject very fully, with explanations of practical results in many adjudicated cases. It is believed that the following instance of the application of the law in recent years will have some weight in a discussion of proposed changes in our mining laws. The litigation between the St. Louis Mining & Milling Co., and the Montana Mining Co., Ltd., a British corporation, over portions of the Drumlummon lode covered more than 20 years, and resulted in a judgment for damages in favor of the St. Louis Co. of such an amount that the Montana Co. declined to pay the judgment, and therefore lost the entire property. The Drumlummon lode claim in the Marysville district, Mont., was located in 1876 by Thomas Cruse. When he staked out his claim, he assumed that the discovery on the Drumlummon vein, and another showing of mineral about, 200 ft. distant, were on the same vein, and the center line of the claim was given a course of N. 579" E. Development of the Drumlummon vein soon showed that its strike was about N. 15" E., and that the other outcrop was on the North Star lode; so it was inevitable that the Drumlummon vein would cross one or both side lines of the location, and the Montana Co. found it advisable to acquire adjoining claims on both sides of the Drumlummon, in order to avoid possible litigation over apex rights. They bought the Marble Heart, but would not purchase the Hopeful (see Pigs. 1 and 2), at the terms proposed for the reason that exploration in the Cruse level had shown the vein to be barren in that region. The owners of the Hopeful were sinking an incline from that claim in 1889, and the Montana Co. enjoined them, contending

Jan 1, 1915

By G. Y. Chin, M. T. Dolan, W. L. Mammel

The least work analysis of constrained deformation, first proposed by Taylor and recently modified by Chin, Hosford, and Mendorf to include twinning, has been applied to the problem of axisynmetric flow. Calculations were made, for the case of (111}(110) slip and {111}(112) twinning applicable to fcc deformation. The solutions were obtained by linear programming techniques and with the aid of computer methods. The results show that in axisymnzetric tension the onset of twinning tends to decrease the plastic strength anisotropy as compared with the case of defortnation by slip alone. In compression, however, the strength anisotropy is increased as a result of twinning. The [100] and 11111 orientations and their vicinity are unfavorable for twinning in tension and compression, respecticely. With slight modifications, the present results may be applied to bcc metals deform ing by {110}(111j slip and (112} (111) twinning. Implications of present results to wire texture development of low stacking fault energy fcc materials arc discussed. PREVIOUSLY we' have applied linear programming techniques to obtain solutions to Taylor's least work analysis2 of constrained deformation of single crystals undergoing axisymmetric flow, such as occurs in an axially loaded polycrystalline aggregate. These solutions yielded values of the parameter M = o/.r = Y/w which relates the external stress-strain (u-c) curve to the shear stress-shear strain (7-y) curve for the crystal. The value of M is dependent on crystal orientation and hence enters significantly in texture strengthening considerations.3'4 Recently the Taylor hypothesis was modified by Chin, Hosford, and Mendorf' to include mechanical twinning. The latter has now been applied to the problem of axisymmetric flow and the results are reported in this paper. Calculations have been made for {111 )( 110) slip and {111 )(112) twinning applicable to fcc crystals. With minor changes in the indices of the slip and twin systems, the present results are also applicable to bcc crystals which slip on (110) (111) and twin on {112)(111) systems.

Jan 1, 1970

By J. L. Shafer, Christopher L. Caenepeel, Martha L. White

Often an in situ leach is the only practical economic method for copper recovery from small low grade oxide deposits. The decision to develop a copper property by an in situ blast and leach is strongly dependent on ore grade, tonnage, hydrology, and the copper extraction and acid consumption rates. The inherent leachability of the ore will in part determine the desired amount of fracturing of the ore deposit by the blast in order to obtain a copper extraction rate that is economical. For the last several years Occidental Research Corp. (ORC), has undertaken a modest effort to address the question of how to best identify and determine the information necessary to predict and control the operation of an in situ leach for the economic recovery of copper. Modelling of the Leaching Process During the past 10 years the activity in the mathematical modelling of the leaching of copper sulfides in copper dumps and heaps has increased dramatically. The modelling of copper oxide leaching has not received as much attention. The acid leaching of non-sulfide copper minerals while chemically less complex, compared to sulfides, does generate a vertical acid concentration gradient in the ore heap being leached due to the acid-base nature of the reaction. This is generally not observed in sulfide deposits that are biologically active, due to the buffering action of opposing reactions involving the generation of sulfuric acid from sulfide minerals and acid consuming host rock. Thus, the objective of ORC 's copper oxide modelling efforts is to determine the leaching parameters which will account for this vertical acid concentration gradient. Knowledge of this gradient will result in the correct choice of the feed acid concentration to the top of the ore dump or heap, such that the effluent will have the desired pH.

Jan 1, 1980

By T. Egleston

In the year 1881 I presented to the Institute two papers containing the resnlts of researches on gold, in which I endeavored to make plain some of the causes of the losses in the working of gold-ores. Since that time I have continued this investigation at intervals with the following results. In the papers previously presented to the Institute I made no attempt to make compounds containing more than two substances. Since then I have endeavored to make more complex compounds. The experiments consist in fusing mispickel free from gold with a certain amount of metallic gold, and experimenting with that; also fusing the same mineral with sulphide of antimony, and fusing iron pyrites with gold and compounds of antimony. The mixture was made with mispickel taken from a vein not far from Stamford, Conn., absolutely free from gold. It was fused

Jan 1, 1884

By Forrest K. Pence

Bentonite deposits are known to occur in Texas within the Jackson group of formations. This group represents the uppermost Eocene age sediments found in the coastal plain area of Texas. It outcrops across this area of the state in a narrow band of some 4 to 20 miles width. The outcrop pattern roughly parallels the present Gulf of Mexico shore line and is some 100 miles inland from the Texas shore, Fig 1. The principal bentonite deposits are found in the areas where this outcrop pattern cuts across the south-central Texas counties of Karnes, Gonzales, and Fayette. In these deposits, the better quality bentonite is found in the lower or bottom layers of the volcanic ash deposits in which they occur. Some of these better quality benton-ite~ develop very light colors upon firing and therefore justify their being classified as "white firing." The deposits in Karnes and Gonzales Counties apparently occur in commercial quantity, whereas the white firing strata so far uncovered in Fayette County have been too thin to be classified as yet as "commercial." A study of the ceramic properties of the weathered ash in Gonzales and Karnes Counties was reported in 1941.' Commercial development of the deposit in Gonzales County, 7 miles east of Gonzales, Texas. was started earlier by the Max B. Miller Co. for the purpose of marketing the material as a bleaching clay, and this operation has developed to very sizable proportions. In recent years, this company has offered a specially selected grade of the Gonzales material as a suspending agent in glaze slips. The white firing property especially adapts the material to use in white cover coat enamels. The strata in the deposit are practically horizontal and consist from top to bottom of approximately 2 ft of soil overburden, 10 ft of brown bentonite, 2 ft of coarse white bentonite, and 4 ft of waxy white bentonite overlying a he grained sandstone. The & being made in the quarry is approximately one-half mile in length. Only the bottom 4 ft of waxy bentonite is being recovered, the upper layers being stripped and wasted, Fig 2. It may appear somewhat surprising that the very bottom strata appears to have been the one most completely altered. To confirm this, samples from top to bottom of the various strata were studied microscopically by R. F. Shurtz. Professor of Ceramic Engineering, University of Texas. His interpretation is to the effect that the lower part of the seam was deposited at a much earlier date than the top, and that the lower part was chemically altered to a considerable extent before the upper part of the seam was laid down. The conclusion to be derived from these examinations may be stated briefly to he that the alteration in these strata or parts of strata has proceeded independently of the alteration in other parts of the strata during a considerable geological period. The presence of gypsum and iron stain throughout all of the strata indicates that alteration is now proceeding more or less uniformly throughout. It is contended that the alteration of the original ash to montmorillonite is not a result of the presently operating processes. A deposit which occurs approximately 7 miles southeast of Falls City and just south of the village of Casta-howa, has been explored and leased by J. R. Martin, of San Antonio. Mr. Martin has conducted mining and marketing operations in bentonite for a period of many years and asserts that the white firing strata found in this deposit occurs in commercial quantities. His pit, which is shown in Fig 3, exposes 2 ft of soil overburden, approximately 5 ft of white bentonite having coarse texture, and approximately 5 ft of waxy white bentonite which in turn overlies a brown sandy clay. Here, as in the Gonzales deposit, the most completely altered portion is found at the bottom of the seam, as per following report of microscopic examination by Mr. Shurtz. Sample No. 1: This sample was taken from the top stratum which is one foot thick. It is grayish in color and it contains visible fossilized plants. The color is probably the result of fine carbonaceous material in the rock. Under the microscope the sample is seen to consist of glass and feldspar; the amount of glass predominating. Both these substances are slightly altered. No montmorillonite or other clay mineral can be identified definitely; however, the products of the slight alteration mentioned are probably montmorillonite or mineral gel. Sample No. 2: This sample was taken from the stratum second from the top. This stratum is fourteen inches thick. The sample is light gray. It shows numerous veinlets of greenish translucent material ranging from one-eighth inches wide down to the limit of visibility with the unaided eye. It has the smooth, sub-conchoidal fracture characteristic of some bentonites. Microscopically the sample consists mainly of aggregates of clay minerals. The birefringence of the aggregates is lower than would be expected if the

Jan 1, 1950

By J. W. Whitlow, P. K. Theobald, W. C. Overstreet

Monazite placers in a region of the Carolinas ex- in 1951-1954 by the USGS are estimated to contain at least 53,000 short tons of thorium and 4600 short tons of uranium. None of these deposits is being mined. The 5200-sq mile area investigated extends northeast along the western Piedmont from the Savannah River in South Carolina to the Catawba River in North Carolina. Much of the region is 700 to 1100 ft above sea level. Gentle hills and broad interfluves give a local relief of 100 to 200 ft. This increases to 1700 ft at the west margin of the area where the east flank of the Blue Ridge is approached in South Carolina and the South Mts. are entered in North Carolina. This report is restricted to placers in the Piedmont. Monazite deposits in the Blue Ridge and coastal plain were not covered by the investigation. Monazite Mining in the Carolina Piedmont: Discovered in 1879 by W. E. Hidden,' the monazite placers in the Carolina Piedmont were first mined in 1886. The area was a major world source for thorium and the lanthanides until 1895, when Brazilian monazite entered the trade. Total output from the Carolina Piedmont has been nearly 5500 tons of monazite,2 but the placers have been idle since 1911, except for 48 tons produced between 1915 and 1917. Monazite mining was revived in the Carolinas in 1954 when a dredge operated by the Heavy Minerals CO. opened a fluviatile placer along Horse Creek in the coastal plain near Aiken, S. C. Geology and Mineralogy: Two recent papers",' have described the geology of the monazite placer district in the Piedmont between the Savannah and Catawba rivers; the general geology of the region has been reviewed by P. B. King.' These reports show that the placers are in locally derived alluvial sediments that cover the floors of shallow, narrow valleys in an area underlain by monazite-

Jan 1, 1960

By Robert K. Steunenberg, Irving Johnson, James B. Knighton

The activity coefficient of plutonium in liquid magnesium, over the temperature range 650° to 800°C, was obtained from measurements of the distribution of plutoninm between a 50 mole pct MgC12-30 mole pct NaCl-20 mole pct KC1 molten-salt mixture and liquid Zn-Mg alloys. For dilute solutions (0.08 at. pct Pu) the activity coefficient of plutonium was found to vary from 10.1 at 650°C to 12.2 at 800°C. The activity coefficients of plutonium in dilute liquid solutions of plutonium in uranium, silver, lanthanum. cerium, and calcium were estimated to be. The distribution data indicate a value of about 0.1 at 800°C for the activity coefficient of PuCl3 dissolved in the above ternary salt mixture. LIQUID magnesium and several liquid alloys of magnesium with metals such as zinc and cadmium have been shown to be useful solvents in pyrochemical processes for the recovery of uranium and plutonium from discharged nuclear fuels,' and for the separation of transuranium elements.' The present study was undertaken to determine the activity coefficient of plutonium in liquid Pu-Mg alloys in support of process-development work. The activity coefficient of plutonium in liquid magnesium was determined from experimental data on the distribution of plutonium between a liquid ternary MgC12-NaC1-KC1 salt mixture and various liquid Zn-Mg alloys. The distribution data were used to calculate the ratio of the activity coefficients of plutonium in liquid zinc and in liquid magnesium. The activity coefficient of plutonium in liquid magnesium was then computed from the known activity coefficient of plutonium in liquid zinc. It was not necessary to know the thermodynamic properties of the molten-salt system explicitly. The major features of the Pu-Mg system have been reported by Schonfeld.3 At the temperatures of interest in the present study, i.e., above about 600°C, the phase diagram indicates the existence of a wide liquid-miscibility gap, with the plutonium-rich liquid containing about 8 at. pct Mg and the magnesium-rich liquid containing about 10 at. pct Pu at the intersection with the solidus regions. Additional data on the compositions of the two equilibrium liquid phases obtained in this laboratory4 have defined the miscibility gap up to the consolute temperature (at about 1040°C). EXPERIMENTAL PROCEDURE AND RESULTS Materials. The 50 mole pct MgC12-30 mole pct NaC1-20 mole pct KC1 salt mixture was prepared by melting the required proportions of reagent-grade NaCl and KC1 with anhydrous MgC12. The molten salt was then purified by contacting it with liquid Cd-30 wt pct Mg alloy (at 450°C) to reduce oxidizing impurities, followed by filtration through a stainless-steel frit (pore size, 65 µ) to remove solid MgO formed during the reduction. The purity specifications of the zinc, magnesium, and plutonium were 99.999, 99.8, and 99.85 pct, respectively. Apparatus. The liquid salt and metal were contained in a tantalum crucible inside a graphite secondary vessel. The crucible assembly was located inside a resistance-heated stainless-steel furnace tube. The furnace tube was closed by means of a stainless-steel cover, which was attached by bolts, with a neoprene O-ring serving as a gas-tight seal. The top of the furnace tube was water-cooled to protect the O-ring. The furnace-tube cover was provided with a tantalum thermowell, a tantalum stirrer, and a port through which sampling tubes could be inserted and materials could be added to the melt without admitting air to the furnace tube. Vacuum and an argon atmosphere were available through a side-arm on the furnace tube. The furnace temperature was regulated by a proportional controller that was actuated by a chromel-alumel thermocouple between the furnace tube and the heating elements of the furnace. The melt temperature was measured by means of a chromel-alumel thermocouple in the tantalum thermowell. The accuracy of temperature measurement was ±3°C. The salt and metal phases were intermixed by a motor-driven tantalum paddle positioned at the liquid interface. The tantalum crucible was provided with four baffles to increase the turbulence. The sampling tubes consisted of 1/4-in.-OD tantalum tubing that terminated in a tantalum frit (Kawecki Chemical Co.; average pore size, 30 µ). Procedure. The zinc, magnesium, plutonium, and salt were charged to the tantalum crucible; then the system was evacuated and filled with argon. The melt was brought to the desired temperature, and agitated for 1 to 2 hr. After allowing the salt and metal to separate, both phases were sampled. Filtered samples were obtained by immersing the end of the sampling tube in the liquid and increasing the argon pressure sufficiently to force the liquid salt or metal through the frit into the tantalum tube. The sample was then partially withdrawn into the cooler portion of the furnace tube and permitted to solidify before being removed. The temperature sequence for sampling at each magnesium concentration was 800°, 700°, 600°, 650°, and 750°C. The composition of the liquid-metal phase was varied by incremental additions of magnesium in a series of experiments at low magnesium

Jan 1, 1967

By W. Joseph Schlitt

The types of copper leaching and recovery operations in the United States are reviewed, along with characteristics and trends in the industry. The contribution of the various leach operations to total mine output is discussed in order to provide a perspective on the relative importance of copper leaching as a method of production.

Jan 1, 1980