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By M. D. Hassialis, C. G. Myer

THE nature of a collector-coated mineral surface has been the subject of some experimentation and much speculation. Various aspects of the problem have been isolated and studied; it is probable, however, that there exist aspects as yet unrecognized, which are needed for a complete description of the system. The recognized elements of the problem are: 1—The nature of the chemical units comprising the film and the nature of the forces holding these chemical units to the mineral surface; 2—The orientation of the chemical units at the surface; and 3—The disposition of the collector units at the mineral surface. This report touches upon all three aspects of the broad problem; it deals, however, with a new and heretofore unrecognized element—the mobility of the collector units in the two dimensions of the solid-liquid interface. Regardless of the viewpoint adopted as to whether the chemical units comprising the collector film are collector ions or molecules, held either by chemical or physical forces, the writers on flotation theory are in apparent agreement as to the orientation of the chemical units at the mineral surface. Wark and Cox' showed that the contact angle at a stable collector-coated surface in a constant chemical environment is independent of the nature of the mineral and of any resurfacing agent; it depends solely upon the collector used. They showed further that the value of the contact angle increases with the number of carbon atoms in a homologous series of compounds. Taggart, Taylor, and Knoll2 postulated that the collector ion is oriented with the chemically reactive, polar end toward the mineral surface and the non-polar, hydrocarbon-like end toward the liquid phase. In support thereof they cited the abstraction of p-p1 dihydroxydiphenylthiourea and of glycol xanthate8 by galena without a concomitant increase in the contact angle above that of clean galena. It was argued that in these cases the water-repellent layer which normally would be presented to the liquid phase owing to the orientation of the collector ions is overlain by a layer of water-avid groups, the hydroxyl groups. Implicit is the analogous orientation exhibited by a monomolecular layer of an insoluble substance at an air-water interface. It should be pointed out that these facts also are consistent with the concept that the collector units be flat on the surface of the mineral and that the response of the surface to an air bubble is determined by the ratio of the area covered by water-avid groups to the area covered by water-repellent groups; the areas being weighted according to the relative watef-avidity or water-repellency of the groups involved. The disposition of the chemical units of the collector film at the mineral surface has not received the attention it merits. Taggart4 concluded from the observed stoichiometry of the ions entering and leaving a surface that the stoichiometry is obeyed in detail at the surface of the mineral. Thus, in the case of xanthate-coated galena, pairs of xanthate ions are attached to surface lead ions. In the case of chemisorption, Gaudin and Preller postulated an agreement between the atoms or ions of the substratum (the mineral surface) with those of the superstratum (collector film); in the case of the physical adsorption their position is not clear. In one place they state "the first type (physical adsorption) does not require any agreement between the atoms or ions of the substratum with that of the superstratum" and in another place that "It seems, in fact, as though some slight influence of the substratum is formed even in physical adsorption." The present work shows that in some cases of chemisorption the chemical units of the collector film are in constant surface-bound motion and may not be assigned to any site of the surface, hence there is no permanent association between the chemical units of the film and the atoms or ions of the solid. The physical property of the collector film that is of paramount importance is its surface energy. Although this property is not susceptible to direct experimental determination, the related contact angle is. Much of the knowledge of the properties of a collector film derives (by deductive and inductive reasoning) from the behavior of an air bubble thereat

Jan 1, 1952

By Laurance Reid

HIGH compression ratios, resulting from high injection pressures and relatively low recovery process pressures, constitute a major problem, which has confronted those engaged in gas recycling for condensate production since the beginning of that practice. Despite the numerous process developments and improvements, no entirely satisfactory solution for this prob-lem has been presented to date. Often, where the residue gas from the recovery process must be recompressed for subsurface injection, there is a substantial difference between the required injection pressure and the plant-operating pressure required to produce maximum recovery of condensate. These limitations result in high compression ratios, which, in turn, neces-sitate excessive compressor investment and operating cost. As an alternative, the operating pressure may be increased in order to lessen the compressor requirements but such a saving often is purchased at the expense of decreased condensate recovery, caused by operation at the higher pressure. Regardless of the choice, the net result is a diminished return on the capital invested in the project. The goal of future development in this field must be the complete recovery of all valuable hydrocarbon components in the reservoir fluid at pressures that, regarded as compressor suction pressure, will require minimum compressor horsepower to inject the residue gas into the formation. A com-pression ratio of 1.5 has been established as an optimum for modern compressor equipment." Therefore, in order to con-form to this optimum condition, a 3000-lb. injection pressure would require that the condensate recovery process be operated at 2000 lb. per sq. inch.

Jan 1, 1940

By D. H. Davis, I. Fatt

Oil bearing rocks, usually found at depths of 2,000 to 10,000 it, are elastically deformed by overburden pressure. Although the change in porosity with pressure in a number of example rock.: has been shown to be small, it was thought that overburden pressure might have a significant effect on permeability. This note presents results of an investigation to determine the magnitude of the change in permeabilty with pressure. In routine core analysis permeabilities are measured on rock samples which are not under overburden pressure. If permeablity measured in this way differs from permeabilities measured under overburden pressure, a systematic error is introduced into well productivity calculations. The experiments described in this report were made to determine the magnitude of the change and thus to give an estimate of the error introduced into the calculations. The error introduced by the neglect of the change in permeability when overburden pressure is removed may account for part of the difference between well productivity calculated from core analysis and the actual well productivity. The results obtained in this investigation have only quali-tative significance because of the difficulty in reproducing in the laboratory the stress conditions on the rock buried in the earth. The assumption which is usually made is that the rocks in the earth are under a uniform pressure equal to the weight per unit area of the vertical overburden column from the rock in question to the surface of the earth less the pressure of the liquid in the pores of the rock. For an overburden consisting of sandstones and shales (of assumed average specific gravity of 2.3) and for a liquid pressure equal to the pressure in a salt water column (of specific gravity 1.0) reaching from the rock to the surface, the resulting net pressure on the rock is approximately 0.56 psi per ft of depth. EXPERIMENTAL METHOD Clean, dry sandstone core plugs one in. in diameter and about three in. long were mounted in a copper foil jacket as shown in Fig. 3 or molded in a Lucite jacket as shown in Fig. 4. All cores except those labeled "2" in Fig. 1 and "8" in Fig. 2 were measured in the copper foil jacket. Cores "2" and "8" were measured in the Lucite jacket. The jacketed core was placed in a high pressure hydraulic bomb in which the liquid pressure outside the jacket could be raised to 15,000 psi. Flow lines from the core were brought out of the bomb through special fittings in the bomh head and connected to a laboratory type gas permeameter. Measurements were made with nitrogen gas; the maximum mean gas pressure in the core was 16.1 psia. RESULTS The specific permeability of sandstone decreases with in. crease in overburden pressure. Most of the decrease takes place over the range of zero to 3,000 psi overburden pressure. If the net pressure on the rock is assumed to be 0.56 psi per ft of depth as discussed above, 3,000 psi of overburden pressure is equivalent to about 5,000 ft of overburden. At 3,000 psi overburden pressure, the permeability of the eight sandstone core samples tested ranged from 59 to 89 per cent of the permeability without overburden pressure. Figs. 1 and 2 show the permeability reduction as a function of overburden pressure. The permeability of the eight cores without overburden pressure ranged from 4.35 md to 632 md; these permeabilities are listed in Figs. 1 and 2. REFERENCES 1. Carpenter and Spencer: "Measurements of Compressibility of Consolidated Oil Bearing Sands," U. S. Bureau of Mines, R.I. 3450, (1940). * * *

Jan 1, 1952

By H. S. Martin

THE Phosphate Recovery Corporation operates three flotation plants, Nos. 1 and 2 concentrators about three miles northeast of Mulberry, Florida, and No. 3 plant at Wales, Tennessee. These plants represent the first commercial phosphate flotation operation and are also believed to be the largest nonmetallic flotation operation in the world at the pres-ent time. The details of this paper will be confined to plant No. 2, this being the newest and largest concentrator in Florida, and the flow sheet being practically the same as at plant No. 1. The No. 3 plant has been oper-ated but a short time and has been shut down since December, 1932, on account of the decrease in phosphate rock production. The phosphate washers usually are located centrally on a known deposit so that about two hundred acres of ground can be mined eco-nomically to each washer location. The washers are moved to new areas when the distance becomes too great for economical pumping. The flotation plants are built either adjacent to a washer, so that current debris may be utilized, or on the site of a former washer where there is sufficient accumulated debris available for mining, and the flotation plants must also be moved from time to time. Most of the water used in the Florida flotation plants is recovered after passing through a circulating system where most of the fine sus-pended matter is settled out, but some make-up water is obtained from deep wells near the plant. Power is obtained from the Tampa Electric Co. through the dis-tribution facilities of the International Agricultural Corporation. Flotation concentrate is hauled from the plant by trolley over the International Agricultural Corporation's trolley system, to the drying plant at Prairie, one mile northwest of Mulberry.

Jan 1, 1933

By David C. Grant, Erich W. Tiepel

In situ mining is a process for recovering uranium from relatively small or low grade ore bodies. The process involves passing a lixiviant and oxidant, typically NH4HC03 and H202, through a uraniumcontaining ore body. As the solution progresses through the ore body, the uranium is oxidized from the plus four to the soluble plus six state. The bicarbonate complexed uranium laden solution is removed from the ore body and the uranium recovered in an above-ground processing plant. During the leach process lixiviant cations may be sorbed on the ore host, if the host contains exchangeable clay minerals. This report examines the implications of this clay interaction on ground-water quality, the various available in situ mining restoration methods, laboratory procedures for simulation and development of those methods, and the results of using the restoration methods in the field. Various restoration methods, including ground-water sweeping, clean water recycle, and chemical restoration are presented and discussed. The use of ore characterization, agitation leach tests, and column tests to study cation sorption and desorption, and thus, to identify and evaluate restoration methods is discussed. Ore characteristics identify the clay type and clay content, which allows the estimation of the magnitude of the cation sorption by that ore. The chemical effects of clay exchange and of chemical restoration can be examined using agitation leach experiments and packed ore column tests which also provide information on the physical effects. The agitation leach test is a relatively rapid means of determining the sorption and desorption characteristics of an ore. Unlike agitation leach tests, column tests can be used to assess the influence of solutions on permeability, on precipitation problems, and on other chemical changes. The effectiveness of using restoration methods, which were developed and tested in the laboratory, are examined by presenting field restoration data from several mining sites.

Jan 1, 1979

By V. Lessels, D. J. Buckwalter, D. H. Rose

A cyanide leaching process developed at Michigan Technological University was tested in a 12 to 24 tpd pilot plant. Flowsheet layout and equipment are described. Cyclic operation of leaching, copper precipitation by acidification, and regeneration of the cyanide leaching solution are discussed. Copper extraction and reagent consumption were as predicted by laboratory tests, and the process was proven to be technically feasible. The White Pine Copper Co., a wholly-owned subsidiary of the Copper Range Co., produces copper in the Upper Peninsula of Michigan. Production in 1965 amounted to 138 million pounds of metal. An expansion program now under way will increase this by approximately one half. The highly mechanized room and pillar mine produces 18,000 tons of ore per day, which is treated in a flotation plant. The concentrate is smelted in a reverberatory furnace, the matte is blown in converters to blister copper, and the blister copper is further refined to a uniformly high-quality Lake copper, which is 100% conductive and has a silver content of 25 to 40 oz per ton. The run of the mine ore averages 22.5 lbs of copper per ton with the ore mineral occurring in the form of chalcocite. The nature of the mineralization is such that the sand tailings average a little over 4 lbs of copper per ton. Intensive research has improved copper recovery in the flotation operation to some extent, but at the scale of present operation about 12,000 tons of sands containing over 48,000 lbs of copper go to the tailings pond daily. When the plant expansion is completed, the loss of copper will be somewhat greater. In 1959, the White Pine Copper CO, initiated a study of the recovery of copper in the sand tailings. Inasmuch as the flotation operation was considered to be optimized, a logical approach seemed to be leaching. For various reasons ammonia, ferric chloride, acid, and biological leaching were ruled out as impracticable. This led to a consideration of cyanide leach-

Jan 1, 1968

Albert SaUveur, Cambridge, Mass. (communication to the Secretary*):—On close examination I think it will be found that the evidence by which Mr. Corson claims to have shown the inaccuracy of a statement I made a few years ago, is far from convincing. This statement is: "Hot work as such has no influence upon the structure of the metal. Indirectly, however, by retarding crystallization until a lower temperature is reached, it may influence its structure most decidedly ; but the same results could be accomplished by heat treatment alone, i.e., by re-heating the unworked metal to the temperature from which the worked piece was allowed to cool undisturbedly." The Melallographist, vol. ii., p. 267 (under the head of "Changes of Structure Brought About by Work"). Of the two pieces of steel tested by Mr. Corson, one was worked and finished at a a cherry-red" and the other reheated to a cherry-red, and the assumption was made that in both cases the temperature was about 715' C., although no pyrometric device was used to record it. The possibility of considerable difference in temperature alone is so great as to invalidate Mr. Corson's inferences. The apparent temperature,

Jan 1, 1907

The Ortiz Mountain property is located approximately 40 km (25 miles) southeast of Santa Fe and 13 km (8 miles) from the village of Cerrillos. The property embodies a Spanish mineral grant from which Gold Fields acquired about 2, 200 hectares (900 acres of surface and mineral rights. Following a period of exploration, metallurgical testing and development of a feasibility study in 1974, the decision was made, because of the higher gold price in the fall of 1978, to place the property into production. the gold is confined in a quartzite breccia; it is an unusual occurrence because the gold in this host rock is leachable at minus 19 mm (3/4 in. ). Total ore reserves are estimated at approximately 3.6 million mt (4 million st) averaging 1.89 g per mt (0.055 oz per st) based on a cut-off of 0.7 g (0.02 oz). The ore will be recovered by open pit mining at a stripping ratio of 1. 2:l. Production is scheduled for 13,600 mt (15,000 st) of ore per week or approximately 680,000 mt (750,000 st) per year. Primary crushing of pit-run material is by a 1,070 by 1,015 mm (42 in. by 40 in.) jaw crusher. Crushed ore is conveyed to stockpile and then to a 2.13 m (7 ft) standard crusher with the screened over- size crushed in a 1.5 m (5 ft) tertiary cone crusher. The screen undersize, the final product for leaching, is minus 19 mm (3/4 in. ). A 1, 360 kg (3,000 lb) sample is cut from the crushed ore stream, further crushed in a small jaw crusher and then sampled in two more stages by Vezin samplers to produce a final 13.6 kg (30 lb) sample for assaying. Pebble lime is then added to the ore on the belt as it is sent to the conveyor gantry spanning the leach pads. A tripper on this moving gantry stacks the leach ore without compaction. The entire crushing plant and conveying system is of open-air construction. Leach pads have a total length of 515 m by 60 m (1,700 ft by 200 ft) wide with usable dimensions of 490 m by 58 m (1,600 ft by 190 ft). The directional length slope is 3. 870 with 370 on the width. The pad is constructed of 175 mm (7 in. ) thick asphalt containing a rubberized asphalt membrane. It is divided by a 300 mm (1 ft) high curb into eight segments. Each segment is loaded with 13,600 mt (15,000 st) of ore, which is one week's production. Optimum working height of a pad is considered to be 3. 0 m (10 ft) with a maximum height of 4. 0 m (14 ft).

Jan 1, 1981

By S. P. Sharples

Having occasion a few weeks ago to investigate the subject of black band iron ore in West Virginia, I was surprised at being met at the outset with a denial of its existence. One gentleman went. so far as to offer one hundred dollars an acre for any land in West Virginia that contained a seam or vein of even a single foot in thickness, and when shown the ore he acknowledged that it was genuine, but declared it had been imported. That black band does occur in West Virginia is to-day well known to a number of persons. There is a seam on the Gauley River of about a foot in thickness of good ore, which I understand Mr. Lewis and others are making preparations to work. The great trouble with black band ores as heretofore found in this country has been their limited extent. At Fort Washington, for instance, they had a very fine show ing of ore, but it soon gave out. It has also been asserted that the black band ores of the Western Coal Measures were worthless in the furnace, and Fort Washington has been pointed out as an example of the fate of any one who should undertake to work them. But they have been successfully worked at Youngstown and in other parts of' Ohio, and I am assured by old furnace-men from Ohio that there is not the slightest difficulty in. working them if properly managed. The deposits to which I wish to call the attention of the Institute,

Jan 1, 1882

By S. P. Sharples

HAVING occasion a few weeks ago to investigate the subject of black band iron ore in West Virginia, I was surprised at being met at the outset with a denial of its existence. One gentleman went so far as to offer one hundred dollars an acre for any land in West Virginia that contained a seam or vein of even a single foot in thickness, and when shown the ore he acknowledged that it was genuine, but declared it had been imported. That black band does occur in West Virginia is to-day well known to a number of persons. There is a seam on the Gauley River of about a foot in thickness of good ore, which I understand Mr. Lewis and others are making preparations to work. The great trouble with black band ores as heretofore found in this country has been their limited extent. At Fort Washington, for instance, they had a very fine show ing of ore, but it soon gave out. It has also been asserted that the black band ores of the Western Coal Measures were worthless in the furnace, and Fort Washington has been pointed out as an example of the fate of any one who should undertake to work them. But they have been successfully worked at Youngstown and in other parts of Ohio, and I am assured by old furnace-men from Ohio that there is not the slightest difficulty it working them if properly managed. The deposits to which I wish to call the attention of the Institute,

Jan 1, 1882

By C. E. Silverblatt, D. A. Dahlstrom

Previous bench-scale studies demonstrated that a substantial moisture reduction can be obtained by the application of steam to coal filter-cakes, particularly to those containing a limited amount of minus 200-mesh material. Consequently, U.S. Steel's Applied Research Laboratory conducted a study to determine the feasibility of steam filtration of 1/4 in. by 0 coal on a continuous pilot-scale installation in which an Eimco extractor is used. The test results showed that the moisture content of this coal could be reduced to 5.5% at a steam loading of about 175 lb of steam per ton of dry cake. In terms of steam consumption, this loading is calculated to be 1-1/2 lb of steam per lb of water removed. In some tests, only 1 lb of steam per lb of water removed was required. These levels of steam consumption are significantly less than the actual heat requirements of conventional thermal dryers. The satisfactory moisture reductions were obtained at filtration rates of 125 to 186 lb of dry solids per sq ft per hour. However, the proper design of commercial filters should improve the potential rates that can be attained. These results indicate that steam filtration has considerable economic potential for drying 1/4 in. by 0 coals and also has additional advantages over the conventional thermal drying of coal. The results of the previous bench-scale studies and those published by Burton' demonstrated that a substantial moisture-content reduction can be effected by the application of slightly superheated steam to coal filter-cakes, particularly to those containing a limited amount of minus 200-mesh material. Consequently, US Steel's Applied Research Laboratory conducted a study to determine the feasibility of steam filtration of 1/4 in. by 0 coal on a continuous basis. In the continuous process, the dewatering operation on the filter is conducted in stages. In the first, the bulk of the water in the feed slurry is removed by conventional mechanical dewatering on a portion of the filter before the steam is applied. In the second stage, steam is applied to the cake. The steam condenses and releases its latent heat, thereby reducing the viscosity of the remaining cake water; this reduced viscosity then facilitates further drainage and produces a cake of lower moisture content. Because of the availability of feed coal on a continuous basis and since steam was readily available from a low-pressure boiler, the Applied Research Laboratory selected for the test site a coal-preparation plant that cleans Pittsburgh Seam coal. A slurry of table and flotation coal in percentages as they occur normally was used as feed stock. Typical analyses of this slurry are shown in Table I. DESCRIPTION OF FILTER AND EXPERIMENTAL PROCEDURES An Eimco Extractor was selected for the evaluation of steam filtration in a continuous process. This unit was chosen because of its availability and its design features that permit relatively simple hood construction, instrumentation, and operation. A photograph of the Extractor is shown in Fig. 1. In the filter a cloth medium is used that travels horizontally and continuously with a grooved, rubber drainage belt. The total effective filter area was 12.5 sq ft. As the feed slurry flows by gravity from the feed box onto the medium,

Jan 1, 1964

By David D. Rabb

Leaching or solution mining, a relatively simple and economical process for beneficiating metallic ores, is likely to find increasing application in the treatment of low-grade ores that are impractical to mine by any other means. This process may be carried out in two different ways: 1) dump leaching, where the ore is moved from its original location to be leached at another site; and 2) In-situ leaching, where the ore is leached in place by introducing the leach solution at the top, letting it flow down through the ore under gravity, and then recovering it plus the dissolved metals it contains. Whichever leaching method is used, it is almost always necessary to break up the ore before leaching. In this paper a study is reported which indicates that rock broken by an explosion-in particular, an underground nuclear explosion-is significantly more amenable to leaching then is rock broken by other methods. These results suggest that the leaching speed and efficiency could be increased by nuclear fracturing of the ore. Not only would the leach time be shortened, but the resulting increase in strength or richness of the solutions would decrease plant installation expense as well as reduce pumping and processing costs. A considerable fund of experience has been accumulated in the course of several hundred experimental underground nuclear explosions, so that the gross results of any given nuclear explosion can now be predicted with a fair degree of confidence.' From this knowledge it seems clear that, under the proper conditions, large ore bodies can be fractured much more economically-macroscopically speaking-by nuclear explosions than by other methods. The present study concentrates on smaller scale effects that is, the cracks in the chunks of rock broken by the explosion-and shows that here too, in the microscopic domain, there are important advantages to nuclear fracturing. The intense shock produced by the very fast acting, high-brisance nuclear explosive fractures the rock in a way that should significantly improve its leachability. Experimental Procedure This study compared rocks broken by nuclear explosives with rocks produced by conventional mining, quarrying, or core drilling. The test samples, granite chunks 6 to 8 in. on a side, plus core sections, came from the area of the Hardhat*2 nuclear explosion and were taken both before and after the explosion. For comparison, several samples of quarried granite were obtained from a local gravestone monument company. The general procedure was to soak the test samples in leaching solution and then determine the extent of penetration. A standard commercial copper leaching solution was used (10 gpl Cu, 10 gpl H2SO4, 5 gpl ferric Fe, 15 gpl total Fe, pH about 1.5), to which a water-soluble penetrant dye, Zyglo 1-c, had been added. Details of the procedure were as follows: 1) Sample leached in solution containing Zyglo penetrant dye. 2) Washed with water. 3) Air-dried. 4) Cut with granite wire saw. 5) One face polished with granite monument polish. 6) Sent directly to be photographed, or heated at 110°C for 2 hr and then sent to be photographed. 7) Photographed under ultraviolet light to show crack patterns. Results After 10 days of leaching at 70-75°F, the samples were removed from the solution, washed, dried, and cut in half with a granite wire saw to study the penetration of the leach solution. Since the Zyglo dye in the leach is visible under ultraviolet light, the degree of penetration of the leach (and hence the cracks in the samples) can be studied on photographs of the crosscut samples made under ultraviolet light. The photos in [Fig. 1] show how the leach solution penetrated various representative samples. Of the 71 rock samples examined, fractures were most frequent and prominent in samples from the rubble produced by the nuclear explosion [(Fig. 1D)]. Fracturing was less apparent in shaft-mined rock [(Fig. 1B)], still less evident in drift-mined rock [(Fig. 1C)], and practically nonexistent in cored or quarried specimens [(Fig. 1A)]. The samples in [Fig. lA-C] were from the same general area as the nuclear explosion, but they were obtained before the explosion. Results of the crack studies are summarized in [Table 1]. The Zyglo-treated leach solution penetrated the test samples at the rate of about 1/2 mm during the first hour, 1 mm by the end of 4 hr, 2 to 3 mm in 12 hr, and 4 to 6 mm in 10 days, showing a progressively slower rate with time.

Jan 1, 1972

By R. R. Sayers

TWO serious hazards from coal dust confront the bituminous-coal miner- -a physical or safety hazard and a physiological or health hazard. The first threatens the miner with loss of life from coal-dint explosions, and the second with loss of health and earning capacity from inhaling large quantities of the dust. Studies by the Bureau of Mines have shown that coal dust, except some grades of anthracite dust, is capable of initiating an explosion when no explosive gas is present and that its property of extending or propagating an explosion has been responsible for nearly all the widespread explosions that have occurred in bituminous mines in the United States in the present century. For years many mining people believed that low-volatile coal dust was not explosive but not only is it explosive but if the source of ignition is intense, it is as explosive as the dust of many high-volatile coals because the war is creating increased pressure for greater coal production and because of the trend toward greater mechanization in mines, explosion and health hazards from coal dust are likely !n increase considerably.

Jan 1, 1943

GENERAL GEOLOGICAL DESCRIPTION The uranium mineral appraisal herein described was made for an ore deposit in New Mexico. The uranium mineralization occurs as roughly tabular to lens-shaped deposits of varying thickness within the sandstone of the Jackpile formation of Jurassic age. The mineralization occurs at several horizons within the Jackpile formation, but is not always confined to specific beds. It often transgresses across stratigraphic layers, as demonstrated when attempting correlation of uranium gamma anomalies between adjacent drill holes. Preferred mineral trends have been established in some parts of the deposit, but are not always predictable when attempting to extrapolate beyond the zones defined by drill hole sampling. MANAGEMENT OF DRILL HOLE SAMPLE DATA All of the drill hole sample data used in the appraisal were obtained from interpretation of gamma-ray logs. A brief description of gamma-ray log sampling is given in Chapter 2. Two preliminary steps are involved in preparing gamma-ray logs into sample data normally acceptable for computer manipulation. 1) The gamma analog trace produced by the downhole probe must be digitized, i.e., counts-per-second (CPS) are recorded for each 0.15 m (1/2-ft) interval across the gamma-ray anomaly. The work is usually done on a "digitizer" while following the analog trace with an electronic stylus on a special tablet or table. The other alternative is to do the work manually. Some of the newest downhole logging equipment produces a magnetic tape output which can be read by a computer. 2) The gamma-ray CPS values must then be converted to equivalent U308 assay values, taking into consideration all appropriate correction factors. The US Atomic Energy Commission, now part of the Department of Energy, developed the theory and algorithm for this conversion in the 1960's, and though the work can be done manually, it is best done with the GAMRED (gamma-reduction) computer program. Fig. 14-1 illustrates a typical analog trace produced by the probe instrumentation, including gamma-ray trace (left), self-potential electric log (center), and resistivity log (right). Fig. 14-2 shows a form of the computer output after the analog data has been digitized to 0.15-m (1h-ft) intervals and converted to percentage U308 values for the same hole. The U308 grade bar graph on the right permits rapid visual inspection of anomalous values by depth and elevation. The converted U308 sample data were next processed by the SURVEY program that assigned to each assay interval [0. 15-m (1h-ft)] an X Y Z coordinate, taking into account all deviations indicated from downhole survey information. The output file of the SURVEY program was listed on the line printer for examination by project geologists, and also became the main input file for the URAN/ COMP compositing program, which is designed to group individual 0.15-m (/-ft) U308 sample values into larger intervals representing discrete anomalous zones according to a series of specified parameters entered into the program by the user. Composite U308 intervals were computed by the URAN/COMP computer program, the logic of which may be briefly described as follows: Step 1 The program searches for the highest U308 value in the hole and builds a composite by adding the highest grade interval on either side so long as the added value is greater than a minimum acceptable grade, and the total composite, after addition, meets minimum GT (grade times thickness) constraints. Step 2 The program loops back to find the next highest assay value in the hole not already included into a composite, until all anomalies above minimum cutoff grade are composited. Step 3 Individual composites are combined if they are closer than a specified minimum waste thickness, and the resultant composite grade still meets GT minimums.

Jan 1, 1980

By A. L. Holley

The existence of iron ore and anthracite coal in the neighborhood of Providence, R. I., baa long been known, chiefly as a geological fact; that these materials, so near to each other and to tidewater, are of such a good quality and are present in such large quantity, as to have seriously raised the question of establishing blast furnaces there, was a surprising fact to me; and I have thought that the few notes I have lately gathered on the subject would be of interest at this partly New England meeting. The coal field referred to has an area of above 400 square miles, and is found throughout the belt of transition rocks extending from Newport Neck to Mansfield, Massachusetts. It underlies the cities of Providence and Newport, and the towns of Middletown, Portsmouth, Jamestown, Warwick, Barrington, Cranston, North Provi-

Jan 1, 1879

By Robert Sisselman

Central Sweden's Grängesberg underground iron ore mine, which accounts for more than three million tons of pellet product annually, is experiencing a major changeover to continuous block-caving. The newly-adopted method, complete with a total revamping of mining equipment, will be fully operational by 1975. For the past 15 years, Sweden's Grängesberg mine has been mined using two different methods. Those ore thicknesses of 117 ft or greater (they comprise about 70% of the total area) are worked by block-caving techniques with cross-cutting carried out on each 195-ft working level, and loading at the foot wall on each 58-78-ft level. Sub-level caving is carried out in all ore sections less than 117 ft thick.

Jan 1, 1974

By Frederick Prime

THE Great or Cumberland Valley, which (under a variety of names) extends from Canada, through Vermont, Massachusetts, Connecticut, New York, Pennsylvania, Maryland, Virginia, and East Tennessee, to Alabama, is composed in part of limestones lying immediately above the Potsdam sandstone. Geologically, they belong to the era between the Potsdam and Trenton, and are commonly termed Calciferous. A peculiarity of these limestones is their great richness in limonite or brown hematite ores, which form a large part of the charges of the many blast-furnaces situated on their outcrops. The mines from which the ores are obtained are (properly speaking) only pits or quarries, from which the ore is extracted by means of picks or shovels, powder being rarely employed. The ore does not occur casually at any point in the limestone, but forms regular lines, following apparently the 'outcrops of certain beds. In Lehigh County, Penn., where I have been engaged on the Geological Survey of' the State, during the past season, there are three or four of these lines. During the progress of the survey I was much struck by the fact that the two most important and promising lines of outcrop were, one at the base of the crystalline schists (forming the South or Lehigh Mountain), and overlying the Potsdam sandstone conformably; the other, near the line of junction of the calciferous limestone with the slates of the Cincinnati epoch. The mines along these two lines, following the topography of the country, were in place, richer and more permanent than those of the belts between, which were generally leaner. A peculiarity of these two lines of outcrop, and one to which I particularly desire to draw attention, is the occurrence of what I at first mistook for a highly altered slaty limestone, but which subsequent analysis proved to consist in great part of damourite (hydrated potash mica). The following are analyses of this damourite slate I. From Fogelsville, Lehigh County, within a few hundred feet of the contact of limestone and slate, by Dr. F. A. Genth. II. From Hensingerville, Lehigh County, within 300 feet of gneiss, by Mr. Sydney Castle, of the University of Pennsylvania. III. From Allentown, Lehigh County, within 150 feet of con-

Jan 1, 1875

By Peter F. Weiss

INTRODUCTION Austro-American Magnesite Co. of Radenthein, Carinthia, Austria, is a wholly owned subsidiary of General Refractories Company, Bala Cynwyd, USA, and has been incorporated since 1908. The raw material basis is the magnesite mine on Millstatt Alp, located in the mountains near Radenthein at an altitude of 1500 m above sea level. The mined crude magnesite is crushed in a surface crushing plant to minus 35 mm and consequently hauled to the processing plant in the valley by means of a cable conveyor of 7.5 km length. The annual production of 350,000 mt is accomplished with 20 supervisors and 122 workers (of those 60 are in the mine). The labor requirement for block caving is 0.2 man hr/mt, corresponding to an output of 40 mt/man/ shift. GEOLOGY The rock in this area mainly consists of crystalline type schist. There are mostly coarse crystalline garnet mica schists with varying mineral concentrations, and to a lesser degree, some deposits of amphibolite and marble. This petrographically rather homogeneous rock is divided into numerous different tectonic units (covers and scales). The magnesite deposit itself is a foreign body that also has been tectonically inserted. The ore body has a length of 600 m, is 450 m deep, and has a thickness of 60 m on average, varying substantially however, The contact with the accessory rock is very precise. The deposit 1 = ore body, 2 = previous open pit, 3 = caved in area, 4 = block under excavation, 5 = land slide consists of coarse crystalline magnesite in excess of 90 per cent. The few impurities consist of dolomite and talc. The lower part of the deposit is inclined parallel with the schists at approx. 40 degrees. In profile it can be seen, however, that the upper part of the deposit has been tilted over. This is a result of a post-glacial land slide reaching from the bottom of the valley at 1400 m above sea level to the peak of the mountain at 2100 m above sea level. Due to this land slide, with horizontal movements of up to 100 meters, as well as older tectonic events, the rock was crushed to a vast extent. Through mining, the land slide was reactivated; about 40 million cubic meters of rock slowly flow downhill almost like a plasticized mass. Near the surface, the rock is completely crushed and resembles a glacier rupture zone. This renders the originally impermeable rock completely water permeable so that practically all of the annual quantity of 1000 to 1100 mm of precipitation permeates the rock and flows down into the mine. The carbonate rock turns into the preferred aquifer. From 1700 to 2600 cubic meters of water have to be pumped from the mine daily. On one hand, additional problems are caused by the water in the mine, on the other hand we have learned to make use of the down- hill pressure forces to improve the crushing process during block caving. The boundary of the downhill shearing zone moves downward as mining goes deeper.

Jan 1, 1981

By Effingham Humphrey

THE Lehigh Valley. Coal Co. finished the rebuilding of its Drifton No. 2 breaker at Drifton, Pa., in the summer of 1917. The new construction comprises an addition and the complete remodeling of the old breaker. It is of interest to note that the old structure was the first iron breaker erected in the anthracite region. THE OLD BREAKER This old breaker was built by the Cross Creek Coal Co., under the direction of the late Eckley B. Cole, in 1888-89, and was fully described by him in the Transactions.1 The columns were cast iron, 8,'10 and 12 in. square. The floors and partitions in pockets were cast-iron ribbed plates. The struts were cast iron of "H" section; the other members were of structural iron either rolled I-beams or built-up sections. The sizes of the rolled beams varied from 6 to 20 in. deep, while the built-up sections, made of plates and angles, varied from 6 to 63 in. in depth, so that there apparently was no rule to determine the type of section used. In the remodeling of the old structure, practically everything was taken clown excepting the columns and roof, and such beams as were necessary for temporary construction purposes, and, therefore, a close inspection of the old structural members could be made and the effect of the acid water oil them could be noted. The cast-iron members were but slightly attacked by the water, while the wrought-iron beams, in the wet places, were corroded almost to failure.

Jan 2, 1918

By P. Eadington

I feel sure that the apparent anomolies between his findings and ours arise as a result of the different experimental methods used. The modem oxygen electrode method he used is sensitive to nano moles of oxygen uptake whereas the method we used was only sensitive to micro moles at best. Furthermore, our experimental method did not permit measurements to be made in the first ten to fifteen minutes of reaction, the very time he is observing. Direct comparison of his data to ours is made difficult because of not knowing the specific area of the samples he used. However, we found that precipitation of PbS using sodium sulphide gave rise to samples of area approximately 30 m2 g - 1. Noticing the rapid decline in oxidation rate over ten minutes, as reported in Table 1, one can conclude that we could not measure the rate after ten minutes using our method. Taking the above area I estimate that at these rates only a small fraction of surface would be oxidised in a period of hours. I can only conclude that at some point a change in oxidation mechanisms occurs giving rise to a much faster rate such as we observed. Finally, I would like to comment on the apparent inertness of the homogeneously precipitated lead sulphide sample 11. This method of preparation we found to give specific areas of 20 to 30 times less than other methods and this would account for the very low oxidation rate.

Jan 1, 1982