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By E. F. Thomas

THE late George S. Rice was active in the inves--I- tigation of bumps, particularly in the last ten years of his career as chief mining engineer of the U. S. Bureau of Mines. Since most of his investigation was carried out in Great Britain, continental Europe, and—to a lesser extent—Canada, his thinking on prevention was influenced considerably by the experience of those countries. It is not surprising, therefore, that when he was called upon a few years before his retirement to investigate bumps in the U. S. and suggest ways to prevent them, he turned to longwall mining. A longwall method had been most successful in combating the bump hazard in mining coal under deep cover, especially in Great Britain, but the prevailing method there at the time was advancing longwall mining, which he knew was uneconomical under U. S. mining conditions. For this reason he proposed a modified retreating longwall system that he believed included the best features of the advancing method. As brought out by Rice,' if the cover is 2000 ft and 50 pct of the coal is extracted, the static load on the remaining pillars will be about 4000 psi, which exceeds the ultimate crushing strength in most instances. If the pillar coal is overloaded before a pillar line is established, then the abutment zone preceding a line of extraction is no place to split pillars or extract them by any method other than an open-end system. Rice therefore advocated open-end mining, preferably by longwall, but he was willing to compromise with long-face mining if the longwall method was not acceptable. Rice's system was put into operation in a mine in Harlan County, Kentucky,3 but subsequent experience has shown that it did not take into account two important factors—avoidance of pillar-line points and maintenance of adequate development in advance of the pillar-line abutment area. For ten years after Rice's retirement the USBM did little investigation and research on bumps, chiefly because so few were occurring that there was not much cause for alarm. But in 1951 there were three occurrences involving fatal injuries, and the Bureau began a statistical survey in that year. C. T. Holland, head of the department of mines at Virginia Polytechnic Institute, was retained as a consultant. The resulting study' of 117 case histories brought out these important conclusions: 1) Almost invariably the bump occurred in a locality affected by the abutment zones of one or more pillar lines. 2) In most cases the locality of the bump was influenced by the abutment zones of more than one pillar line. The term pillar-line point has been used for many years in the Appalachian region for such a situation. Point is used in the geographical rather than the mathematical sense. 3) In pillar-line extraction the following practices are safest in preventing bumps: a. The mine layout should provide for pillars of uniform size and shape along the extraction line. b. The mine layout should be planned so that no development need be done in the abutment zone of a pillar line. c. The layout should permit open-end extraction of pillar lines from the next goaf, so that it will not be necessary to resort to pocket mining, splitting pillars, or any practice that will involve driving in the direction of the goaf within the abutment zone. d. Pillars should be large enough to support area without undue roof and floor convergence before establishment of a pillar line. These are, of course, generalities, and while they are useful in laying out areas where bumps can be expected, they are of limited help in many mines that were committed to a system of mining before it was realized that they were subject to bumps. Under such conditions it becomes necessary to choose between the following alternatives: 1) Abandon the territory, except for pillars that offer no extraction problems. 2) Through experience select the pillars that are most heavily loaded, and, by augering, induce bumps from a safe vantage point so that impinged loads are relieved. This method was first developed at the Gary, W. Va. mines of U. S. Steel Corp. and later adapted to mining thick coal beds at Kaiser Steel's Sunnyside mine in Utah. No scientific method is available to determine where to drill within a loaded pillar. Although this method of unloading has worked very successfully at Gary—with one exception—

Jan 1, 1959

By E. Kendall Cork

The traditional financial objective for a single mine company has been to operate as frugally as possible and to pay out most of the earnings as dividends. If the business is cyclical (as it is for most metals) the dividends might fluctuate quite widely. When the mine is exhausted the company disappears. This is still quite a viable strategy for a single mine company. It is not however a viable strategy for the world as a whole. The mining industry is built by mine development companies who can mobilize the people and capital to bring new mines into production. Their skills must include marketing, engineering, finance and other politics. It is very rare for a property to be brought in without the support of a major company that can provide all these services. The exceptions will usually have some other form of big brother support, for example the U.S. government uranium contracts at guaranteed generous prices. The mine development company will seek as a minimum to perpetuate itself by developing new mines in order to replace those which are running out. The more common and more ambitious objective is to grow -- that is to add to its ore reserves and current production by developing more new mines. The financial objectives for that company are very different. Obviously if all the earnings were paid out in dividends there would be nothing left to work with. The first financial policy then is to spend an appropriate amount on exploration for new properties. The next is to retain enough of the earnings to provide the capital for new projects at least sufficient for the equity. There is no magic formula as to what proportion of earnings should properly be distributed as dividends by a growth-oriented mine development company. As a rough rule of thumb distributing half or more will probably leave too little to work on and 30% or so is probably a good balance. However the circumstances differ widely from company to company. It may be useful to set an objective for the rate of growth of a company's earnings. Some have picked rates such as 15% per annum compounded. Others have set a target in real terms which might appear as 10 or 11% plus inflation. Obviously the arithmetic of compound interest is very attractive; however in practice there is much variation. Indeed current returns from existing operations swing widely with the business cycle and there is no assurance that economic new properties will be found according to someone's arbitrary time schedule. For example, Western Mining Corporation Limited in Australia explored for 30 years with little to show for it, but then found the great Australian nickel deposits and more recently the huge Roxby Downs copper. That long dry spell could not have fitted anyone's arbitrary calendar of growth and yet they would not have found such orebodies without that long period of effort. Should they have abandoned the search? Once a new property has been found or acquired there has to be a threshold rate of return on the new capital to be invested against which to evaluate the property's economics. Conventionally this seems to be 15% after tax, a number common in other heavy industries as well. In some cases it is expressed as a lower number plus allowance for inflation. Discounted cash flow analysis is a very useful tool but it does not make the decision. In the end a "go" decision depends on judgment of many factors some of which are numbers used in the DCF calculation whose credibility must be examined. It is curious how frequently investment proposals come in with the rates of return very close to 15%. The project advocates know that a number much less than 15% will not fly and that a number much more is not necessary. With much higher nominal and real interest rates of recent years, even though before tax, logic suggests that the hurdle rate should also rise. The power of compound interest is so great that 20% is very hard to achieve in any cash flow projection but 18% may be a sensible yard - stick. Once again it is remarkable how many project proposals come in with an 18% return. On the record the mining industry as a whole has not been overly restrictive in choosing its hurdle rates of return. This is shown by the abundance of metals in recent years and the failure of metal prices to keep up with inflation. All of the foregoing is standard text book stuff.

Jan 1, 1985

By R. A. Garling

INTRODUCTION Since the infancy of in situ uranium mining, a growing number of hydrometallurgical processes have been incorporated into pilot and commercial scale flowsheets. Although initial design efforts were geared toward maximizing uranium recovery and minimizing plant and wellfield flow circuit maintenance, recent emphasis has shifted to improved means of water conservation and aquifer restoration. As mining units approached depletion, evaporation ponds reached minimum freeboard, and state and federal agencies demanded proof of groundwater restoration, processes including mixed bed and conventional ion exchange, reverse osmosis and electrodialysis were adopted by the industry. These units served the additional function of reducing process bleed flows during mining in states where the deep disposal well permitting ice remains unbroken. This report concerns the use of electrodialysis as an alternative to the more conventional processes used in in situ mining. In addition to a brief history and description of the process, a comparison to reverse osmosis and operational data derived from testing an Ionics, Inc. 1.31 x 10-3 m /s (30,000 gallon/day) unit at the Teton-Nedco Leuenberger Research and Development pilot will be presented. HISTORY Commercially practicable electrodialysis was contingent upon the development of synthetic ion exchange membranes in 1940's. In 1952, Ionics Inc. demonstrated that the process was amenable to the treatment of salt and brackish water and, in 1954, made their first commercial sale. The following decade saw several major electrodialysis unit sales which were generally targeted for use on private or municipal potable water treatment. Major increases in membrane desalting unit capacities, facilitated by technological advances in the reserve osmosis industry, were noted during the 1970's. The development of polarity reversing electrodialysis equipment which reduced feed pretreatment requirements, increased water recovery rates, and simplified unit operation, kept Ionics Inc. competetive in the water treatment industry. Engineering advances which incorporated automated equipment, non-corrosive construction materials, and improved ion exchange membranes allowed the electrodialysis process to compete in industrial waste treatment among other commercial markets. PROCESS AND APPARATUS DESCRIPTION The electrodialysis process utilizes direct electrical current passed across a stack of alternating cation and anion selective membranes in order to achieve an electrochemical separation of ionized materials in an aqueous solution. The membrane stack has the appearance of a plate and frame filter press and auxilliary equipment includes solution pumps, electrically actuated valves, filters, piping and a direct current power source. The ion separation membranes are thin sheets of synthetic cation or anion selective resins. Attaching sulfonate or quaternary ammonium groups to the cross linked copolymer structure determines the ion selectivity of the membrane. The membranes are separated from each other in the stack by non-conductive spacers that house flow channels which route the flow tortuously and parallel to the membranes. Direct electrical current passing perpendicularly to the membranes and solution passages attracts cations toward the cathode and anions toward the anode (Figure 1). As the ions from the feed stream pass through the ion selective membranes, they become concentrated in the adjacent brine channel and are retained there by the combined attractive force of the electrode and the repelling force of the next membrane toward the electrode. Limiting factors on the degree of demineralization possible include chemical solubilities in the brine flow and the current density that will produce an unacceptable degree of polarization (Figure 1). Feed or brine solution treatment with complexing agents or acids has been successfully applied to prevent membrane scaling. Polarization can occur when sufficient current density is applied to dissociate water in the ion depleted region of the diluting compartments near the membrane surfaces. Significant polarization is evidenced by large electrical resistances across cell pairs and notable pH differences between diluting and concentrating streams. Limiting current densities have been increased in U.S. manufactured equipment by utilizing tortuous flow paths of relatively high linear velocities thereby promoting continous solution mixing. Energy consumption is due to separating electrolytes and solutions, oxidation and reduction reactions occurring in electrode compartments, overcoming electrical resistance, conversion from AC to DC power, solution pumping and auxiliary equipment actuation. A major improvement to the basic electrodialysis process was applied in 1970 which resulted in frequent, automatic cleaning and descaling of membrane surfaces. The process, polarity reversal, incorporates alternating the cathode and anode on a periodic basis while exchanging product and brine flow channels via electrically actuated values. The reversal reduces the potential of stack plugging with CaCO3 (calcite), CaSO4 (gypsum), and colloidal materials and, in most waters, eliminates feed pre-treatment requirements. For approximately two minutes during and following the reversal, off spec. water is flushed to waste or reintroduced to the feed supply. The usual feed treatment on polarity reversing electro-

Jan 1, 1982

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915; when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps: Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible: Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn-drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blasthole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated' by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Such shots were fired from either end .by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1952

By Victor deWolfe

INTRODUCTION The Henderson Mine, located near Empire, Colorado, utilizes a continuous panel caving system to extract ore as one of the world's major producers of molybdenum. Any mine using a caving-by-gravity technique of mining must rely on closely controlled draw of the caved ore. This control is essential to insure proper caving action, to avoid damaging load concentrations of weight and to minimize the dilution of ore with waste material. Henderson is no exception. Draw control is a major factor in all production planning, from long- range plans to short-range and day-to-day ore scheduling. Draw control is reviewed constantly and administered daily in an effort to optimize production efficiency, ore recovery, and cave management. MINING METHOD The cave at Henderson is massive, moving slowly through large panels that are 244 m (800 ft.) wide by 610 m (2,000 ft.) long. Generally two cave areas are drawn at one time. The areas under active draw vary in size but can be as large as 244 m (800 ft.) by 244 m (800 ft. ) containing 400 draw points. Each draw point contains 45,360 mt (50,000 st) on the average and takes about two and one half years to exhaust. A complete panel is worked for seven to ten years. No pillar exists between panels, but rather a buffer zone of broken ore, or "static face," is left in each panel to be drawn with the adjacent, yet-to-be-caved panel in efforts of minimizing dilution of a working area from an exhausted one. (Figure 1) Production drifts are driven on 24.4 m (80 ft.) centers through the ore body. Between the production drifts are funnel-shaped draw bells on 12.2 m (40 ft.) x 24.4 m (80 ft.) centers to receive ore from the cave. Each bell is accessed by two draw points, one from the production drift on either side, thus forming a 12.2 m (40 ft.) x 12.2 m (40 ft.) draw pattern. Extraction of the ore is by rubber-tired, 3.8 m3 (5 yd3) load-haul-dump equipment. The LHDs then tram the ore a maximum of 49 m (160 ft.) to ore passes. Cave initiation and bell development are done from the undercut drifts which are parallel to and 17 m (55 ft.) directly above the production drifts. Longhole rings are drilled and blasted from the undercut drifts to define the bells and establish the undercut for caving. (Figure 2) DRAW CONTROL Since the cave line at Henderson is constantly advancing, it is necessary to be continually initiating new cave at one end while exhausting it at the opposite end. There must exist, therefore, an angle on the ore-waste contact in the broken rock from initiation to exhaustion. The basic concept of draw control is to keep this angle as smooth and even as possible, particularly at the time of exhaustion. If this is achieved, draw points are exhausted more or less in a line, avoiding pockets of remaining ore surrounded by exhausted areas. These pockets would cause spotty ore extraction at the time of exhaustion, increasing the amount of dilution occurring while introducing the potential for significant weight problems in the production area. To arrive at the desired angle on the ore- waste contact, maximum tonnage percentages are assigned to each row of draw points increasing at 10% or 15% increments (depending on cave size and velocity of draw) working away from the cave line. The available tonnage indicated by these percentages is the maximum allowable tonnage to be extracted from each draw point until the available tonnage percent- age is increased. As the cave moves, these percentages increase for each draw point regularly. However, in general the tonnage drawn from each draw point is kept at about 50% of this allowable maximum in order to maintain adequate available tonnage in the cave to sustain production for seven months if cave initiation were to cease. This available tonnage cushion is a safeguard built into the draw control program at Henderson to accommodate fluctuations in the rate of cave advance. When draw points move past the row of 100% tonnage availability, they are drawn past the desired 50% at the same increments per row until exhausted. (Figure 3) To achieve proper draw control, the number of LHD buckets to be taken from each draw point is assigned daily. The actual buckets taken, which may at times deviate from the

Jan 1, 1981

By Brahm Prakash, R. Schuhmann

Chemical conditions for flotation and nonflotation of quartz with oleic acid as collector and barium, calcium, aluminum, iron, and tin as activators were studied using a simple vacuum-flotation technique in glass-stoppered graduates. The detailed study of barium activation led to an interpretation based on ideal Langmuirian chemi-sorption. FLOTATION of quartz is of practical importance as something to be avoided in soap-floating many types of ores. Clean, unactivated quartz is not floated with fatty acids and soaps, such as oleic acid and sodium oleate, in the quantities normally used for flotation. However, data in the literature indicate that almost any multivalent cation will activate quartz if given an opportunity. Thus, a common problem is to prevent activation of quartz by the various inorganic cations inevitably present in flotation pulps. Wark and his coworkers1 have demonstrated the reversibility of the chemical reactions and adsorptions involved in the activation, depression, and collection of the common sulphide minerals. The procedure in much of their work was to bring a mineral surface to equilibrium with solutions of known pH, collector concentration, and activator concentration, and then to test the floatability of the mineral by contact-angle measurement. From the data, graphs were constructed with pH and reagent concentrations as coordinates. These graphs show fields of flotation and fields of nonflotation, separated by narrow transition regions whose locations are shown by so-called contact curves. From the shapes and locations of the contact curves, which roughly separate fields of flotation from fields of nonflotation, a quantitative understanding of the interaction of the reagents with each other and with the minerals often can be deduced. The study of quartz flotation to be described in this paper follows in broad lines the approach of Wark and coworkers. That is, pH, activator concentration, and collector concentration were varied to find equilibrium conditions of flotation and non- flotation, and the results are presented graphically by means of contact curves. However, instead of testing for floatability by measuring the contact angle on a polished surface, a simple vacuum flotation technique was developed and used. Purified oleic acid was the collector and terpineol the frother. Barium activation was studied in some detail, and exploratory studies were made of activation with calcium, aluminum, ferric iron, and stannic tin. Preparation of Materials Quartz: Large lumps of high-grade vein quartz were crushed dry in a cone crusher and rolls. The —20, +28-mesh portion was screened out and used in the subsequent steps. This material was passed through a high-intensity magnetic separator to discard iron, then leached twice with hot concentrated HCl and washed repeatedly with distilled water. The cleaned sand was then wet ground with porcelain balls in a porcelain pebble mill, deslimed repeatedly by settling and decantation to discard —800-mesh material, and again washed with hot HCl followed by distilled water. The resulting stock of quartz was stored under water. Chemical analysis gave 99.8 pct SiO2. Table I gives the size analysis of the quartz used for flotation tests. Calculations from these data, using shape factors given by Gaudin and Hukki9 indicate a specific surface of about 500 cm2 per g. Blank flotation tests in distilled water, and in water with added frother, showed the prepared quartz to be completely nonfloatable and thus indicated the absence of organic contamination. Oleic Acid: The preparation of oleic acid was based on fractional vacuum distillation of methyl oleate2,3 followed by regeneration of oleic acid, and finally fractional crystallization of oleic acid from acetone solutions at low temperatures." The pure oleic acid was stored in a refrigerator. The iodine number of the oleic acid was found to be 90.0 (theoretical 89.93). Oleic acid was used in the form of a dilute water solution of sodium oleate, after preliminary flotation tests showed no effects of form of addition and order of addition of reagents when an adequate conditioning time (that is, 30 min) was provided. Other Reagents: Sodium hydroxide solutions low in carbonate were prepared by first making 1:1

Jan 1, 1951

By J. N. Breston, R. V. Hughes

Conclusions drawn by previous research workers with reSPect to the relation between Pressure gradients and/or velocity and oil recovery obtained by laboratory water flood tests have been in disagreement probably due to variable procedures and unnatural conditions and materials. The Bradford Laboratory of the Pennsylvania Grade Crude Oil Association as part of its secondary recovery research program has conducted nineteen water floods on two long cores of widely differing characteristics in an attempt to clarify this relationship and make it an aid in predicting flooding pressures in the field. Unlike previous research procedures the present experiments were conducted with the aim of duplicating field conditions as closely as possible by using long unextracted consolidated cores, a live crude, and natural brines for both flooding and, connate water content. Also, the pressure gradients and flooding velocities were representative of field conditions where similar sands were being flooded. Eleven floods on one core and eight floods on the other core showed increased recoveries and lower residual oil saturation with increased flood pressure gradients and flood velocities. A marked decrease in recovery was obtained from both cores at very low flood velocities. This pressure versus recovery relationship is shown to hold up to the point of water breakthrough and also up to the 100 and 1 produced water to oil ratio point. INTRODUCTION The possibility of water flooding oil sands was suggested by Carl1 of the Pennsylvania Geological Survey in 1880. It is not known when the practice was tried intentionally for the first time, but its beneficial effects were noted in the annual production rate of the Bradford field as early as 1907. The practice was illegal in Pennsylvania until 1921. Early water floods in the Bradford field usually consisted of shooting or splitting the casing secretly to permit subsurface waters to enter the producing sands under hydrostatic head. As it was noted that the benefits of water flooding seemed to be proportional to the quantity of water dumped into the well many also began to utilize surface sources after the practice became legal. It was probably during the middle 20's before many producers realized that the pressurehead of the water upon the producing sand determined the rate and quantity of water that would enter the sand. Hence, rate and quantity of production appeared to be a direct function of input pressures. By 1927 a few producers had ventured the installation of pressure pumps in order to increase water-input rates and production through the combination of hydrostatic and hydraulic pressures. The adoption of pressure flooding and the "five-spot" drilling pattern in the Bradford field were essentially simultaneous. Water-input pressures in 1930 seldom exceeded 300 p.s.i. at the well head or 1100 p.s.i. at the sand face. Since that time, water-flood producers in the Bradford-Allegany fields have gone to higher and higher pressure until today 600 p.s.i. at the well heads is called a low pressure flood. Many high pressure floods now operate at 1300-1400 p.s.i. at the well head. The limiting and advisable pressure at the sand face has been pronounced as that pressure just under what is required to lift the overburden or to cause formation parting: According to this rule any water flood operation utilizing well-head pressures nearly equal in pounds per sq. in. to 1.1 times the average depth in feet to the top of the producing sand would be considered as a high-pressure flood. Despite the higher pressure trends in the Bradford-Allegany field operations and the results of early laboratory water flooding research, the desirability and benefits of high input pressures are still questioned by many operators, particularly in midwestern water flood operations. The present paper recounts a series of 19 laboratory water floods using two long, consolidated cores of widely differing characteristics saturated with live crude and flooded with oil field brines in an effort to simulate field conditions under • various pressure gradients and flooding velocities. For both cores, higher recoveries and lower residual oil saturations were obtained at higher pressure gradients and flood velocities. This relationship is shown to hold up to the water breakthrough point and also up to the 100 to 1 produced water to oil ratio point.

Jan 1, 1949

By D. E. Light, J. Y. Somnay

The use of alkyl acid phosphates and their alkali salts as collectors for the uranium minerals brannerite and uranothorite was investigated. In particular a detailed flotation study was carried out using isooctyl acid phosphate as a promoter. Tests were conducted on an Elliot Lake ore analyzing 0.11% U 3 O 8, chiefly as brannerite, and approximately 9% pyrite. Pyrite was floated with conventional reagents; the flotation of brannerite from the sulfide tails was studied to determine the optimum pH and pulp density to employ with isooctyl phosphate. 4 uranium recovery of over 90% was possible with a selectivity index of 8.95 when these tails were floated at 17% solids and a pH of 1.7. The effect of various modifying agents was also investigated. Aluminum sulfate, lactic acid and sodium silicate improved the selectivity of collection yielding selectivity indices of 9.5 to 11.5. Ferric chloride was found to be a depressant for brannerite. A Bancroft ore which analyzed 0.13% U308, and contained uranothorite was also amenable to flotation using isooctyl acid phosphate. The composite uranium concentrate obtained assayed 0.41% U308 at a recovery of 95.3% and a ratio of concentration of 3.24. Previous investigations1,2 on the flotation of brannerite, pitchblende, uraninite and/or uranothorite from various Canadian ores showed that fatty acids, including oleic acid and tall oils, petroleum sulfo-nates and alkyl acid phosphates, were suitable collectors for these uranium minerals. Additional work carried out by the Mines Branch in Ottawa on brannerite ores from the Elliot Lake area3 used the tall oil Acintol FA-1, or FA-2, as a collector. A recovery of 92% of the uranium in a concentrate which contained about 55% of the weight and assayed 0.2% U3O8 was obtained from a feed of 0.1% U308 in this latter instance. Eigeles et a1.4 carried out flotation on pitchblende ores and found that the most promising results were obtained with alkyl acid phosphates and their alkali salts as collectors. Within the homologous series of alkyl phosphates, the best flotation characteristics were displayed by isooctyl phosphates. Further, isooctyl phosphate derivatives were much less sensitive to hardness forming salts than oleic acid, enabling all flotation tests to be made with tap water. This paper describes tests employing alkyl acid phosphates for the flotation of ores containing brannerite and uranothorite. The flotation work described herein was conducted in conjunction with the development of a leaching process for Elliot Lake, Ontario, brannerite ores. In this process pyrite in the ore was concentrated by flotation and roasted, providing sulfur dioxide which was used to produce sulfuric acid in situ in a liquid-solid extraction process. A 40% conversion to sulfuric acid was indicated at a leaching temperature of 80°C; this resulted in the extraction of 93% of the total uranium after a contact period of 6 hr. A rotary kiln was used as a leaching vessel. Preconcentration by flotation was studied in conjunction with the above process to reduce the bulk of solids for leaching and, consequently, capital expenditure. The materials and methods used in this study will be described initially. Then the results will be presented in two sections, the first dealing with the brannerite ore, and the second with the uranothorite. Finally an economic analysis of the results on the brannerite material will be made. MATERIALS AND METHODS The two ore samples examined were from uranium-producing mines. One was a brannerite ore, a uraniferous quartz pebble conglomerate from the Elliot Lake area, which contained 0.11% U3O8, 4.3% S and 4.5% total Fe. Mineralogical studies of similar samples indicated the presence of quartz, microcline, sericite, pyrite and minor quantities of rutile, pyrrho-tite and zircon. The other ore was from a biotite-rich pegmatite in the Bancroft area; feldspar, quartz and biotite formed the major gangue components. Uranothorite was the principal uranium mineral; the head assay was 0.13% U 3 O 8. Preparation of Feed: The ores were crushed to —10

Jan 1, 1963

By G. Gutierrez, D. A. Elgillani, M. C. Fuerstenau

The zero-points-of-charge of apatite, calcite, and fluorite are pH 6.4, 10.8, and 10.0, respectively. Scheelite is negatively charged above at least pH 3. In this article, the flotation responses of these minerals in the presence of potassium oleate and sodium silicate are described and compared with electrokinetic data. Colloidal silica appears to be the species principally responsible for calcite depression, while silicate anion is the species responsible for fluorite depression. Additions as high as 1 x 10-3 mole/liter silicate have no effect on the flotation responses of apatite and scheelite. Selective flotation of nonmetallic minerals is difficult to achieve with fatty acids or soaps by themselves. As a result, specific reagents are added to aid these separations, and one of the reagents commonly employed for this purpose is sodium silicate. Flotation separations of various calcium-bearing minerals such as fluorite from calcite1-3 and scheelite from calcite2,4 and apatite,2,5 for example, almost always involve the use of sodium silicate. The mechanisms by which sodium silicate functions as a depressant are still not understood, probably for a number of reasons. For one thing, the dissolution process of sodium silicate is complex, giving rise to a number of ionic and colloidal species.' Moreover, the type and concentration of these species depend on the ratio of Na2O to SiO2, the concentration of sodium silicate, and the pH of the system.' At the present time, it is not known which species, colloidal silica or silicate anion, is responsible for depression. If colloidal silica is the species that is adsorbing, then adsorption must occur by electrostatic attraction between the colloid and the mineral surface. Silicate anion, on the other hand, may adsorb either physically or chemically. The objective of this paper is to determine first the active species of sodium silicate and then the conditions under which this species will adsorb and function as a depressant. EXPERIMENTAL MATERIALS AND METHODS Pure samples of apatite (Durango), calcite (Iceland-spar), fluorite, and scheelite were used in this investigation. Pure potassium oleate was used as collector, while reagent-grade HC1 and KOH were employed for pH adjustment. The sodium silicates used were samples obtained from the Philadelphia Quartz Co. In one series of experiments, various sodium silicates containing different ratios of SiO2 to Na2O were added to calcite systems. These sodium silicates contained SiO2-to-Na2O ratios of 3.75 to 1, 3.22 to 1, 2.40 to 1, and 1.60 to 1. All other experiments were conducted with the sodium silicate containing the ratio of 3.22 to 1, which is the one that is normally used in industry. Flotation experiments were conducted with 21/2-g charges of 48 x 150-mesh material in conductivity water with an apparatus and technique described previously. Electrokinetic experiments were conducted with both a Zeta Meter and streaming potential apparatus. Particle size was 48 x 65 mesh for the streaming potential experiments. EXPERIMENTAL RESULTS The first series of experiments involved flotation of scheelite in the absence and presence of sodium silicate. As shown in Fig. 1, flotation response was not affected with even the relatively high addition of 1 x 10-3 mole per liter sodium silicate. Interestingly though, no flotation was effected below pH 6 with this collector addition. The responses of apatite to flotation under these same conditions are given in Fig. 2. Similarly, no depression was obtained in basic media under these conditions. Similar experiments were conducted with fluorite, and in this case, depression was noted above about pH 11 with 1 x 10-3 mole per liter sodium silicate (Fig. 3). When calcite was floated with these same levels of addition of sodium silicate, essentially no flotation was possible above pH 7 (Fig. 4). The effect of collector addition with a constant addition of sodium

Jan 1, 1969

By R. G. Wheeler, J. W. Goffard

The Larson-Miller parameter was used to correlate time, temperature, and indentation creep of magnesium, aluminum, and some of their alloys. In the temperature range 300" to 450°C, the short-time Meyer hardness of pure magnesium was less than that of the magnesium alloys tested, but for long times the pure magnesium has greater indentation creep resistance. Aluminum (1100 alloy) had 1.5 to 2.5 times more indentation creep resistance than magnesium at 300" and 450oC, respectively. Hardening of aluminum with a dispersion of Al2O3 was effective in the time and temperature ranges studied. New technologies have required the development of new materials and the utilization of the more familiar materials for new and unusual applications. The use of magnesium and aluminum and some of their alloys, because of their desirable nuclear characteristics, light weight, low cost, and ready availability, has been extended to the 300" to 450°C temperature range. In this temperature range the basic consideration of these materials must be their rate of plastic flow rather than offset yield strengths. The indentation testing reported here arose from a need for design data for the load-holding ability of supports made of these materials. Test Procedure—Hardness indents were made with a 0.275-in.-diam quartz indentor and a 10.65-lb load. The indentor was made by fire-polishing a spherical surface on the end of a fused quartz rod. The samples were held at temperature in a graphite crucible controlled to ±2°C. A thermocouple was attached to the sample and test temperatures were recorded. The diameter of the spherical indentation was measured at the end of a test period and the compression stress (Meyer Hardness) was determined by: H___________load__________ m = projected area of indent Samples were 1 in. in diam and at least 1/4 in. thick. It was observed that at the higher temperatures and longer times, the quartz indentor would stick to the magnesium sample. The quartz indentor was, therefore, frequently inspected and fire-polishing repeated when necessary. The area of sticking was always a small fraction of the area of indent and was therefore considered to have an insignificant effect on results. Correlation of Hot-Indentation Test Data with Time-Temperature Parameter—Sherby and Dorn' have correlated creep or tensile data of a' solid solutions of aluminum with a temperature and strain-rate parameter suggested by Zener and Holloman. underwood2 used this parameter to correlate creep properties of some steels with hot hardness, and upon the basis of this correlation a means of obtaining creep properties from short-time (and inexpensive) hot hardness tests has been demonstrated. Since the validity of the correlation of creep properties with a time-temperature parameter and the correlation of creep properties with hot hardness have been shown, it follows that hot hardness may correlate with the time-temperature parameter. The hot-indentation data obtained was expressed as Meyer hardness, and was shown to be time and temperature dependent. Correlation of Meyer hardness, time, and temperature with the parameter was made using the relationship: Hm = Meyer hardness t = time, hours T = absolute temperature, OK K = constant A value for the constant K was calculated by equating In l/t + K/T at different temperatures and times but at the same hardness. The correlation was tested by plotting Hm vs the parameter, In 1/t +K/T. Since materials are being sought which have high hardness at low indentation creep, i.e., a high Meyer hardness for long time at high temperatures, low values of the parameter are ofthe most interest. TEST RESULTS Magnesium—Pure magnesium (99.98 pct) cut from extruded rod was indentation tested perpendicular to the rod axis at temperatures of 300°, 350°, 400°, and 450°C for times ranging from 6 sec to 112 hr. Fig. 1 shows the time dependency of Meyer hardness at the four constant temperatures. Fig. 2 shows the correlation of the Meyer hardness of pure magnesium with the time-temperature parameter using a K of 22,720 in Eq. [I]. At the bottom of Fig. 2, the effect of doubling the time of indentation t2 = 2(t1), on the abscissa for any time is shown graphically. This effect is of constant magnitude. Also shown graphically are the magnitudes of the effects on the

Jan 1, 1960

By R. F. Mehl, R. F. Bunshah

A fast amplifier technique has been developed for the measurement of the rate of propagation of martensite in an Fe-29.5 pct Ni alloy. The time of formation of one plate of martensite is 3x10 sec and the rate of propagation is 3300 ft per sec approximately. IT has been known for some time that the plate-like structural unit of martensite forms from austenite with great rapidity. Wiester1 and Hane-mann, Hofmann, and Wiester took motion-pictures of the transformation as it occurs in a 1.65 pct C steel; they demonstrated that a single plate formed fully in the time interval between successive frames, viz., 1/20 sec, thus setting an upper limit. Forster and Scheil,3 using an Fe-Ni alloy with 29 pct Ni, recorded the sonic characteristics of the process electrically, upon an oscillograph, setting the upper time limit at 0.002 sec. Forster and Scheil,~ measuring the change in electrical resistance in the same alloy upon a cardiograph, set a limit of 0.02 sec. Forster and Scheil5 later, employing the same alloy, improved their technique, reporting an upper limit of 7.10 sec. In studying signals of such short duration, it is an important question whether the frequency response of the electrical system used is high enough compared to frequency of the pulse measured, or, put differently, whether the system is able to reproduce without distortion the signal arising, in this case, from the formation of a single martensite plate. Forster and Scheil (referring only to their last paper) obtained signals of a frequency of 30 kilocycles (hereinafter kc); this was about the frequency response of the equipment used; thus, if the signal had a frequency higher than 30 kc, it would still appear as a signal of frequency 30 kc. All of these results thus provided upper limits only. Recent developments in electronics have made available equipment with very high frequency response, very high sweep-speeds, high gain, etc. The electrical characteristics of such equipment, used in the present study, are given in Table I. Such equipment offers obvious attraction in the study of the rate of propagation of a martensite structural unit—and perhaps of other structural alterations proceeding at a very high rate. This paper reports an attempt to develop a technique employing such equipment to measure the time of propagation of a martensite structural unit and the variation of this with temperature, with the mode of formation—athermal and isothermal—in both polycrystalline and single-crystal samples; and from such measurements to obtain the rate of propagation. As will be seen, the results obtained are useful theoretically. Materials All data presented here are for an Fe-Ni alloy of the following analysis: 29.5 pct Ni, 0.027 pct C, 0.135 pct Mn, 0.094 pct Si, balance Fe. There were several reasons for choosing this alloy: 1—it is substantially the one used by previous investigators; 2—it exhibits both the athermala and the isothermal' mode of formation of martensite, both studied in detail by Machlin and Cohen; 3—the subzero temperatures of transformation in this alloy are experimentally very convenient; 4—it exhibits the "burst phenomenon";" 5—the change in electrical resistance upon the formation of martensite, a decrease, is great, approximately 50 pct.' The polycrystalline specimens were in the form of wires of 0.025 in. diameter; the single crystals were 1x1/4x1/4 in. Experimental Methods Electrical Apparatus: Fig. 1 is a schematic drawing of the electrical circuit used. The principle used in these measurements is the same as that used by Forster and Scheil." A small direct current, about 1 to 2 amp, is passed through the sample. When a martensite plate forms, the resistance of the sample changes and a high frequency signal is generated. This signal is picked up by the probes attached to the sample, fed into the bank of amplifiers and thence to the vertical deflection plates of a cathode-ray oscilloscope. The signal itself triggers the oscilloscope trace which flashes across the tube face and is photographed by means of a 35 mm movie camera at the end of a light-tight hood. The camera has no shutter. As soon as the signal flashes

Jan 1, 1954

By A. W. Sommer, H. H. Kellogg

It is shown that SO3-O2 mixtures react with sphalerite at an appreciable rate ill the temperature range of 361° to 527°C to fornz ZnSO4. The rate of reaction follows a parabolle lax. Oxygen, or O2-SO2 mixtures, have a negligible effect on sphalcrite in the same temperatuee range. PRACTICAL roasting of sphalerite is usually performed at 800°C or higher. The calcine is composed of zinc oxide unless the roaster atmosphere contains relatively large amounts of SO2 and the temperature is held to 800oC or lower, in which case zinc sul-fate may also form. Ong, Wadsworth, and assell' have made a careful kinetic study of sphalerite oxidation under these "normal conditions" and have concluded that the rate of the reaction is controlled by the decomposition of an activated complex adsorbed on the sphalerite surface. They found the rate of oxidation to be small at 700°C. Extrapolation of their data to 400°C would indicate an almost negligible rate at this temperature. In our work sphalerite was reacted with air, mixtures of SO2, O2, and N2, and mixtures of SO3, O2, and N2 in the temperature range 361o to 527OC. Negligible rates of oxidation were found, except for the gas mixtures containing SOs. With this latter gas, the oxidation of finely divided sphalerite was fairly rapid and zinc sulfate was the product. Based on the limited evidence available, the rate of sphalerite oxidation by SO3 is postulated to be controlled by gaseous diffusion through pores or cracks in the zinc-sulfate coating. This evidence for the direct reaction of sphalerite and SO, at low temperature may prove of importance to the understanding of zinc-sulfate formation in the dust-collecting equipment usually associated with zinc roasters. The roaster gas carries off fine dust, much of which may be unreacted sphalerite. The temperature in dust-collecting equipment drops from the roaster temperature (about 800°C) through the range of temperature we studied, to ambient temperature. Such dusts are known to contain far larger amounts of zinc sulfate than the primary calcine. It has been assumed in the past that this sulfate is formed by reaction of ZnO dust with SO3 (or SO, plus 0,) in the partly cooled gases. Our work shows that an alternative possibility exists—the direct reaction of SO3 with ZnS dust. EXPERIMENTAL Reaction rates for sphalerite at 350° to 600°C are relatively small so that it was necessary to use finely divided material with a relatively large surface area in order to obtain measureable amounts of reaction. In the preliminary experiments, pure mineral sphalerite, ground to pass a 325-mesh sieve, and chemically precipitated ZnS (in the form of an impalpable powder) were used. In the quantitative rate experiments the ground mineral sphalerite was processed in an Infrasizer to obtain a product with particle size between 9 and 18 µ. The analysis of this material was 66.7 pct Zn, 0.09 pct Fe, and 32.7 pct S (theoretical sphalerite is 67.1 pct Zn, 32.9 pct S). The sized powder was carefully mixed and divided into l-g samples. The apparatus used for the quantitative rate measurements is shown in Fig. 1. For the preliminary experiments the following change was made: The external catalyst furnace (B, C)* was not used. *Letters refer to Fig._______1 .___- Rather, when catalysis of the SO2 + O2 reaction was desired, the glass beads in basket M were replaced by vanadium-oxide pellets. Procedure—A l-g sample of ZnS was placed on a shallow stainless-steel tray, I, and spread evenly to a depth of about 1mm. The tray was placed in the reaction assembly and the assembly inserted in the cold furnace. The furnace tube was flushed with dry argon and then heated to the desired reaction temperature. The temperature was controlled to ±0.5oC. Gas mixtures (O2 + N2 or O2 + SO2) were prepared from tank gases by means of a mixing device identical to that employed by Darken and Gurry.2 The total flow rate of gas was 205 ml per min to point A of Fig. 1. The accuracy of the individual and

Jan 1, 1960

By W. S. Cavender

The Round Mountain mining district, Nye County, Ne- vada, was discovered in 1906 on claims owned by Lewis D. Gordon. Initial mining operations uncovered gold veins of spectacular richness, and within a few days of discovery, Gordon sold his controlling interest for some $87,000. From this sale emerged the Round Mountain Mining Co., predecessor of Nevada Porphyry Gold Mines, Inc., the latter destined to become the major property owner in the area. Vein mining in the district continued sporadically into the early 1930s, yielding 9.3 Mg (330,000 oz) of gold plus substantial silver credits from approximately 626 kt (690,000 st) of ore. In addition to the lode deposits, the early miners recognized the placer potential in the alluvial fan material accumulated around the west and north sides of Round Mountain itself. Intermittent placer operations were carried out for a number of years, and in the 1940s and 1950s, Round Mountain Gold Dredging Co. worked the placers under a lease from Nevada Porphyry Gold Mines. The last placer operation terminated in 1959 when it, like some of its predecessors, proved uneconomic. Total placer production for the district is estimated at 3657 km (4 million yd) of gravel containing 59 Mg (210,000 oz) of gold and possibly 2.0 to 2.3 Mg (70,000 to 80,000 oz) of silver. Round Mountain is a small hill situated on the east flank of the Toquima Range in central Nevada. The hill is com- posed of relatively flat-lying Tertiary rhyolitic ash flow tuffs, which overlie Paleozoic metasediments and Cretaceous granites. Throughout the surrounding Round Mountain mining district, most of the known gold ores occur in the tuffs, although the metasediments and granites are also mineralized. Mineralization is structurally controlled, principally by a series of northwest-trending shears and shattered zones. Vein, stockwork, and disseminated ores occur, usually containing simple quartz-pyrite-gold mineral assemblages. The gold itself is electrum, having a silver content of 30% to 40%. In September, 1967, Elwood Dietrich, a prospector and mine promoter, obtained a purchase option on the 4452 ha (1 1,000 acres) of mineral rights held at Round Mountain by Nevada Porphyry Gold Mines. The original option had a buy-out price of $1 million and was established through Dietrich's friendship with officers of Nevada Porphyry. In April, 1968, Dietrich conveyed his option to Ordrich Gold Reserves Co., a partnership created by a group of west coast investors, mostly employees of the airline industry. There- after, Ordrich invested considerable funds in trying to test and develop the property, but soon recognized the need to seek financial and technical support from the mining industry. In December, 1968, Dietrich contacted Wayne Cavender, then Regional Geologist, Southwest, for Copper Range Exploration Company (CRX) in Tucson, Arizona, and made a data presentation. Shortly thereafter, Cavender was appointed Manager of Exploration and Chief Geologist for Copper Range Co. (parent company of CRX), New York City, and he asked C. Phillips Purdy, CRX Regional Geologist, Northwest, to make an initial property examination. Purdy's one-week field study took place in March, 1969, and resulted in a recommendation that CRX pursue its investigation of the property. The presence of low-grade gold mineralization in both the alluvial gravels and in the bedrock was verifiable, but the placer was deemed to have the greater immediate economic mining potential. At that time, gold was in the $1.41/g ($40 per oz) price range. Working from Purdy's information, Cavender decided to attempt acquisition of the property, and the first in a long series of negotiations was initiated with Ordrich. Basically, CRX felt that the placer had a promising potential for several reasons, including (I) past operators had recovered free gold but not the gold contained in the pebble fraction of the gravels; (2) past operations appeared to have been ineffectively designed or managed and not costefficient; and (3) the price of gold appeared to be poised for an upward move. Negotiations with Ordrich were prolonged and difficult, with CRX competing against several ma* mining companies, but finally an agreement was reached, effective June I, 1970. Gold was then back to $35. It is believed that, in

Jan 1, 1985

By Hsun Hu, R. S. Cline

The effect of rate of deformation on texture formatiotz has been studied with cube-oriented single crystals of 3 pct Si-Fe, compressed 80 pct at two widely different rates. Compression at a low rate (crosshead speed, 0.1 ipm) produces large orienta-tional changes and coarse deformation bands, whereas little change in orientation is observed, particularly at the surface of the crystal, when compressed at a high rate (impact velocity, -500 ips). Thin plates of mechanical twins are produced at the high rate, but their contribution to the deformation texture is practically negligible. The crystal deformed at a high rate has a lower hardness and a more uniform dislocation structure and is morc resistant to recrystallization. There is also a dilference in the recrystallization texture between the slowly and rapidly compressed crystals. These results indicate that plastic flow is less "turbulent" at high strain rates than at low strain ratcs. THE behavior of metals under high-speed deformation has been the subject of wide study in the past decade. Most of these works were concerned with mechanical properties during testing at high rates of loading, the resulting structures, and the theoretical aspects of the observed phenomena. The present status of knowledge and research activities in this field has been documented in recent publications.17' There is, however, very little information available in the literature concerning the effect of high-speed deformation on the texture formation in metals. The possibility of an effect of rate of deformation on texture is suggested by the microstructural differences observed between normally and very rapidly deformed samples. Aside from the strong tendency for twinning (and phase transformation in certain metals) in high-speed deformation, the slip-line patterns and the dislocation configurations show marked differences as a result of large differences in the strain rates. For instance, in mild steel in static and dynamic compression, Campbell and co-worker3,4 observed that coarse slip occurred on several systems in specimens deformed at normal strain rates, whereas specimens deformed rapidly showed only fine slip on fewer systems, the slip-line pattern being very similar to that in specimens deformed at low temperatures and at normal strain rates. The absence of multiple slip was also noticed by Dieter5 in shock-loaded nickel. The dislocation configuration in an iron shock loaded at 70 kbar pressure was shown by Leslie et al.6 to be similar to that produced by light rolling at low temperature—the dislocation lines were mostly straight and uniformly distributed. These observations suggest that the mode of deformation at high strain rates is notably different from that at low strain rates. Consequently, the nature and the extent of lattice reori-entation during deformation should also be different. This investigation was undertaken to test this idea. MATERIAL AND METHOD The remaining part of a single-crystal ingot of a 3 pct Si-Fe alloy,* prepared for an earlier investi- gation,1 was used in the present studies. Specimens 9/16 in. square and 0.365 in. thick were cut from this ingot with (001) planes parallel to the square surfaces, and the [loo] and 1010] directions parallel to the edges of the specimen, within ±1 deg. These machined crystals were etched to remove distorted metal, and annealed at 1300°C for 24 hr in a purified helium atmosphere. The annealed crystals had a hardness of 161 Dph. The crystal orientation was rechecked with X-ray back-reflection techniques. In high-strain rate experiments the crystal was compressed in a Dynapak machine to approximately 80 pct reduction in thickness (from 0.360 to 0.070 in. with an impact velocity of -500 ips). Thus, the duration of loading was approximately 6x 10-4 sec, and the average strain rate was about 103 sec-1. As a result of this rapid, severe deformation, the specimen became a roughly circular thin disc. The original shape of the crystal, however, could still be recognized by a diffusely outlined square area in the center of the disc. For compression at a low speed, the crystal was deformed in a universal testing machine at a rate of crosshead movement of 0.1 ipm. Compared with

Jan 1, 1965

By D. M. Schuster, E. Scala

The elastic effects occurring in the matrix of a composite reinforced by discontinuous fibers were studied by means of photoelastic techniques. A hirefringent plastic was employed as the matrix material with high-strength a Al2O3 whiskers as the reinforcing fibers. It was found that for whiskers aligned parallel to the tensile load direction: (I) Matrix reinforcement occurred along the length of the whisker to within about 5 diameters of the whisker tip for a length to diameter ratio, L/D, of 40. (2) Significant stress concentrations were created in regions immediately surrounding the tips; the fine as-formed tapered ends exhibited the minimum stress concentration whereas the square-ended whisker produced relatively high values, K r 2.5 for L/D = 40. (3) The axial and radial stress distributions could he determined quantitatively; the stress distribution at the whisker-matrix interface was in general agreement with theoretical calculations. (4) The highest source of stress concentration occurred at points of fracture in whiskers which had ruptured after incorporation into the composite. Whiskers aligned perpendicular to the load direction neither reinforced nor caused appreciahle stress concentrations in the matrix. ThE properties of composites have generally been formulated empirically on the basis of macroscopic tests. There are still many questions concerning the mechanical and chemical interactions occurring at the interface between the reinforcing fiber and its matrix. Continuously reinforced plastics and dispersion-hardened metals represent the two extremes of composite reinforcement. Whether artificially produced or formed in situ, the whisker composite falls between these two extremes and is an example of discontinuous fiber reinforcement. The effect of fiber shape, size, surface condition, and end configuration on the stress distribution in the matrix is important since the presence of stress concentrations, especially in any high-strength, thin-wall structure, could become the cause of catastrophic failure. DOW,' in his theoretical discussion of discontinuous fiber reinforcement, has pointed out that appreciable stress concentrations occur at the tips or ends of the reinforcing fibers. It is the purpose of this study to examine directly, by means of the photoelastic technique, the stresses which occur in the vicinity of a discontinuity or whisker tip and to measure qualitatively and quantitatively the effects of whisker geometry when embedded in a birefringent matrix material. The use of sapphire (a A1,O3) whiskers was decided upon as the reinforcing agent because of their high elastic modulus (E- 60 X 106 psi) and in some cases strengths in excess of one and a half million psi. Sapphire whiskers are of further practical interest since they retain large fractions of their strength up to temperatures approaching the melting point (3720oF).2 The whiskers were grown, by the vapor-phase reaction commonly employed,3 as part of this study to insure a physically uniform supply. Tensile and bend tests were performed to verify that the whiskers were of the expected ultra-high strength. The high-strength whiskers can be identified and selected from the combustion boats by surface perfection, shape, and size; otherwise, strength values can vary by a factor of ten or more. Macroscopic defects are quite common on the surface of large bladed whiskers and the principal experiments were conducted with hexagonal whiskers of about 0.002 in. diameter and lengths ranging from 0.1 to 0.4 in. Upon completion of these two phases of the experiment, several birefringent resins were evaluated. A plastic, developed by Zandman,4 which is supplied by the Budd Co., was found most suitable and was used as the matrix material for the whisker composite throughout the study. It has a sensitivity of 60.5 lb per in. order, calibrated in tension and cures without introducing residual stresses. Young's modulus for this matrix material is 430,000 psi in the fully cured condition. Since good wetting and bonding between the sapphire and the matrix material is essential to the interpretation of the results, a wettability test was performed. The contact angle was measured to be 168 deg where perfect wetting is defined as 180 deg; therefore, the Budd Plastic wets the sapphire

Jan 1, 1964

By R. R. French

Investigations were conducted by the lndiana Geological Survey for some dry methods of bene-ficiating low-grade gypsum ore. Seventy-two batch and continuous flow tests were performed with a roller mill, rod mill, pebble mill, electronic color sorter, electrostatic separator, and an air separator. Approximately 650 size analyses and 550 chemical analyses were performed during the investigation. Batch samples were tested by the Survey, and most of the continuous flow tests were handled by com-mercial laboratories. INTRODUCTION Although many companies producing gypsum have been interested in dry beneficiation, very little information has been published in the past. One of the most comprehensive investigations was reported by the Canadian Department of Mines.5 Other reports of interest have been published by the Canadian Department of Mines,' the South Australian Department of Mines,2 and the U.S. Bureau of Mines.3 Some beneficiating practices, such as hand picking, selective mining, crushing and screening, milling, and air classifying, were already in use in the Canadian industry prior to MacPherson's investigations.5 MacPherson examined the effect of tabling, air separation, screening, flotation, electrostatic separation, calcining, washing, and various types of milling, in his efforts to eliminate dolomite, limestone, and small amounts of silica and clay from the gypsum ore. RAW MATERIAL Gypsum and anhydrite occur within the lower part of the St. Louis Limestone (Mississippian) in southwestern Indiana. Gypsum in single beds 10 or more ft thick occur at Shoals, Martin County; near Freedom, Owen County; near Bloomfield, Greene County; and in the Devonian strata of LaPorte County. Only the Shoals deposit has been exploited commercially. The western edge of the Shoals deposit is contaminated by a continuous bed of shale, about 1 1/2 to 2 ft thick, near the top of the evaporite and by thin continuous strata or irregular masses of carbonate rock. The low-grade material used in these investigations was obtained from the waste pile of the National Gypsum Co.'s plant at Shoals. The waste ore averaged about 67.4% gypsum and was contaminated by various amounts of shale, dolomite, and limestone. All the material had been previously crushed and screened to minus 1 1/4 to plus 3/8 in. X-ray analyses of powdered, sedimentated, gly-colated, and heat-treated samples of the shale showed that it was composed of slightly structurally disordered illite, Fe-rich chlorite, and very finegrained disseminated silt. The clays and silt were partly cemented with carbonate material. Light-gray argillaceous limestone and gray or brown porous dolomite made up most of the carbonate rock contamination. Crushing Characteristics: The rate unweathered gypsum, carbonate rock, and shale reduce in size was determined by crushing handpicked samples in a rod mill and screening at minus 100 mesh (.0059 in.). Fig. 1 graphically represents the data. These data indicate that carbonate rock should be relatively easy to separate from gypsum, but that shale should be difficult to separate from gypsum by differential crushing. Controlled Samples: Eight rod mill tests of hand-picked samples of mixed gypsum and shale and mixed gypsum and carbonate rock were made. Sample weight and crushing weight were standardized, but the scalping screen size was varied from 3 to 10 mesh (.265 to .0787 in.) in order to obtain the optimum purity and recovery. The data obtained from the tests (see Table I) illustrate that the relatively hard carbonate rock allows appreciable beneficiation (13.6 to 18%), but the low strength gypsum and shale allows only minimal separation (3.1 to 7.5%).* Semi control led Samples: During the investigation considerable slaking of the shale was noted when the ore was subjected to alternate wetting and drying. As demonstrated in laboratory tests, as much as 76% shale can be removed by two cycles of wetting and drying in a 48-hour period with subsequent screening at .265 in. Five ore samples, stockpiled and weathered, were screened, rod-mill crushed and rescreened at .157 or .187 in. with excellent results (Table 11).

Jan 1, 1967

By P. S. Kotva

Dislocation substructures in superalloy matrices of varyzng co)npositions have been studied. In general, it has been found that the alloys can be classified into ''high", ''medium", and "low" stacking fault energy classes based on the type of dislocation substructure observed in the matrix and that the substructure can be correlated to the stacking fault energy. The effect of different types of dislocation substructure and dislocation reactions on the intragranulur precipitation of carbide phases has been studied. In a Ni-Cv-Mo-Fe matrix, precipitation of MC carbides in association with stacking faults has been observed. In most superalloys, solid-solution strengthening and precipitation hardening are the chief mechanisms employed to achieve strength. The latter contribution to strength is usually achieved by the precipitation of / in certain wrought alloys. Insufficient attention has been given to the problem of obtaining strength in su-peralloys by controlling precipitation of carbides on imperfections within the matrix. The present work was undertaken to investigate the dislocation substructure in various superalloy matrices, to study the effect of such substructure on subsequent precipitation of carbides in the matrix, and to investigate whether certain modes of precipitation of carbide phases found in austenitic stainless steels2"4'6 would occur in nickel-base alloy matrices with dislocation substructures of the same type as those found in austenitic steels. 1) EXPERIMENTAL TECHNIQUES Five-pound heats of the various alloy compositions reported here were vacuum-cast. The ingots were given light deformation by rolling to break up the as-cast structure and then homogenized for 24 hr. HASTELLOY alloy X (nominal composition: Ni-2OCr-17Fe-8Mo-0.05C) was homogenized at 2150°F and In-cone1 625 (nominal composition: Ni-20Cr-5Fe-8Mo-3.5Cb-0.05C) was homogenized at 2280°F. Fabrication of 0.004-in. sheet was achieved by cold rolling with intermediate annealing treatments being carried out at the same temperature as those used for homogeniza-tion. Each solution anneal was followed by quenching. The aim of this procedure was to redissolve as much of the primary carbide phase as possible. Samples of the 0.004-in. sheet were cut and encapsulated in quartz capsules and then heat-treated in the tube furnaces. Thin foils were prepared using an ethanol-10 pct perchloric acid bath at 32°F and at a voltage of 22 v. A "window" technique was employed. Observations were made on a JEM-7 electron microscope operating at 100 kv. 2) EXPERIMENTAL RESULTS a) Types of Dislocation Substructure. Fig. 1 shows a schematic correlation between stacking fault energy, SFE, and the type of dislocation substructure observed in various matrices of nickel- and cobalt-base alloys. A precise quantitative determination of stacking fault energy is not implied in the figure but the correlation between stacking fault energy and the type of dislocation substructure obtained allows alloys to be divided into three classes in analogy with the classification employed by Swann and ~uttin~' for binary alloys of copper. Class I alloys are associated with a "high" SFE and show a cellular substructure of dislocations as typified by the micrograph of a thin foil of pure nickel deformed 4 pct at room temperature in Fig. 2. With decreasing SFE the tendency toward cell formation is lessened and dislocations tend to be arranged in coplanar groupings. Examples of this class of alloys with "medium" SFE are provided by the mi-crostructure of solution-heat-treated, quenched, and deformed thin foils of HASTELLOY alloy X, "Waspaloy" (prior to any aging), and Inconel 625. Fig. 3 shows a thin-foil micrograph of an alloy of Inconel 625 composition, solution-heat-treated, quenched, and deformed 5 pct at room temperatures. No evidence of any cell structure can be obtained in materials of "medium" stacking fault energy, Fig. 3, even after severe deformation. The stacking fault energy of the alloy shown in Fig. 3 is, however, not low enough to make the dissociation of dislocations visually obvious. As stacking fault energy decreases further, with successive addition of solute in the matrix, there is an increased tendency toward dissociation of dislocations and cross slip becomes progressively more difficult. Eventually, when the stacking fault energy is "low" enough, complete dissociation of dislocations is seen to occur as shown in Figs.

Jan 1, 1969

By Paul G. Shewmon

The migration of slightly solzrhle spherical particles through a solid under the infllrence of a temperature gradient is analylzed for the cases of various transport mechanisms. It is shown that the variation of the velocity of the particles with radius, r, depends on the dominant mechanism of matter transport around or through the inclusion. Thus the velocity varies as r-1 jor surface-dijf1~sion controlled migratzon. is independent of r for volume diffusion in either phase, and varies as if the rate is determined by an interfacial reaction (n is the order of the interfacia1 reaction). These same results hold tor the migration of a cvlinder of length much greater than its radius. and .for other types of potential gradients, e.g., an electrical field. These equations are combined with recent electron-microscopic ohservations to show that the rate of migratzolz of small helium-silled bubbles through copper is determined by surface diffusion of the metal atoms. With these equations and the temperature gradients attainahle in an electron-microscope joil, the dominant transpar/ mechanism (or any migrating pnrtzclos can he determined. AS a result of fission, rare-gas atoms are created in the hot fuel element of a nuclear reactor. These insoluble atoms precipitate to form bubbles of the gas in the fuel. The subsequent migration and coalescence of these bubbles is thought to play a dominant role in the swelling of fuel elements which in turn can limit the fuel-element life. As a result of this practical problem and because of the relative simplicity of the system, workers in several laboratories have imbedded rare-gas atoms in metals with the aid of an accelerator and studied the formation and behavior of the bubbles that form on annealing. Recently Barnes and Mazey have studied the migration of such helium-filled bubbles in copper foils using the electron microscope and the temperature gradient that can be induced in the foil by the electron beam. They found that the smaller bubbles moved faster than the larger ones. It is shown below that, if this is true, the rate-controlling step in their migration is surface diffusion of metal atoms instead of volume diffusion, vapor transport, or an interfacial reaction. The analysis given is in no way limited to gaseous inclusions so the variation of velocity with particle size for any relatively insoluble precipitate particles could be used to obtain information about the relative importance of surface diffusion, volume diffusion, and interfacial reaction in such two-phase systems. VOID MIGRATION IN A TEMPERATURE-GRADIENT ANALYSIS Barnes' studies indicate that helium does not dissolve or diffuse in metals to a measurable extent, so that when the voids move they must do so through the movement of metal atoms from the leading to the trailing sides of the void.' Thus voids might move by surface diffusion, diffusion of vacancies in the surrounding lattice, or vapor transport. We consider first the case of flow by surface diffusion alone. We assume that there is a force, in the sense of irreversible thermodynamics, tending to move the atoms around the bubble in a given direction. We designate this direction as the x axis and the force per atom as Fa. The net rate of flow of atoms from one side of the bubble to the other, under the influence of this force, is the surface flux, Js, times the cross-sectional area available for flow, A,. A, is taken as the circumference times the thickness of the high-diffusivity surface layer 6. Thus where 51 is the atomic volume. For bubbles to advance a distance dx, a volume equal to nr : dx must flow around the void so that The minus sign enters because the bubble moves in the direction opposite to that of the force on the atoms. Equations for volume diffusion and diffusion through the vapor can be obtained in the same manner. For vapor transport we have Ag = and the ratio of the densities of metal atoms in the gas and the solid (pg/pl) enters so that

Jan 1, 1964

By Klaus Detert

Fifteen and 18 pct Ni maraging steel and several binavy and ternary alloys of the iron-rich corner of the Fe-Co-Ni system have been studied. After annealing in the austenite range, these alloys were deformed slightly by cold rolling and subsequently aged. Compavison of experimental measurements at various processing stages indicates that small degrees of deforn~ation (5 to 10 pct reduction in thickness) do not influence hardness and strength or the aging kinetics. Electrical resistivity and ductility are somewhat redu]ced and the coercive force is substantially lower when compared to identical samples which were not deformed. The results are explained in terms of retained austenite which transforms into martensite with the small amount of deformation. RECENTLY there has been increasing interest in maraging steel for applications in electric power generators requiring a high ratio of power to weight.' Fifteen pct Ni maraging steel has been investigated to evaluate its properties and performance for application as high-strength magnetic rotor steel,' and a thorough investigation of the precipitation and transformation processes in this alloy has been conducted in this laboratory.3 The influence of deformation, which preceded the aging treatment, was also studied. It was found that a small deformation in the order of 5 pct has a rather striking effect in reducing the coercive force. Other interesting effects on the properties caused by such deformation have led to the more detailed studies reported herein. EXPERIMENTAL METHOD The methods used to produce the samples from 15 pct Ni maraging steel have been described elsewhere.3 The general procedure of hot forging and cold rolling to sheet has been followed. The composition is included in Table I which also contains the composition of the 18 pct Ni maraging steel grade 250 studied. More data about the material and the processing of the 18 pct Ni maraging steel have been recently reported.4 The experimental alloys of binary and ternary Fe-Ni-Co alloys were made by levitation melting in an argon atmosphere and casting into rod-shaped ingots 7 mm thick weighing 20 g. From these ingots 25-mm-long rods were machined to study the phase transformation by a dilatometer. These rods were then cold-rolled to strips, 2.4 mm thick. Annealing for austenitizing was done in tube furnaces in an argon or helium atmosphere followed by cooling in the water-cooled portion of the tube. Aging was done in a salt bath containing a neutral blend of Alkali nitrates and nitrites. Vickers hardness was measured using a load of 50 kg for 15 sec. Tensile tests were carried out on a hard beam tensile tester at a strain rate of 5 X lo-' per sec using flat samples with a gage length of 37.5 mm and a cross section of 1.25 by 6.25 mm. Estimated sensitivity was 5 K x strain and 0.35 kg per sq mm stress. Magnetic saturation was determined on small cylindrical samples, 2.5 mm in length and diameter, by measuring the force exerted by the gradient of a magnetic field of 1000 oe per cm in a mean field of 11,500 oe. The measured values could be reproduced within *1 pct. Coercive force was measured by a precision-type magnetic field probe in a magnetic coil as described by Foerster.5 The magnetizing field was 1300 oe. The samples were 5 cm long, 1 cm wide, and 2.4 mm thick. he coercive force was measured to within +0.1 oe. Electrical resistivity was measured by using a potentiometer on flat strips with a gage length of 100 mm and a cross section of 2.5 by 0.5 mm. The accuracy of the measurement was 0.1 pct but the reproducibility of the measured specific resistivity among different samples was in the order of 1 pct. The method used to prepare the samples for electron transmission has been reported elsewhere.3 RESULTS Fig. 1 shows the influence of deformation after annealing on the resistivity, hardness, and coercivity. The samples were annealed for 30 min at 1000"C,

Jan 1, 1968

By G. G. Hancock, H. H. Johnson

Hydrogen gas at atmospheric pressure is shown to cause substantial embrittlement in a martettsitic high-stretzgth steel. Subcritical crack growth is observed at very lom stresses and with high growth rates. It is suggested that an adsorbed layer of hydrogen atoms at the crack-tip surface provides the damaging hydrogen. Oxygen in small quantity, as little as O.7 pct by volume, is shown to eliminate subcritical crack growth in gas environments caused by hydrogen or water vapor. However, dissolved oxygen in water has little or no influence upon crack-growth rates. Oxygen-stopped cracks may be restarted by water, or by hydvogen or water vapor if oxygen is removed from the environment. These observations are interpreted in terms of preferential adsorption of oxygen at the crack tip and the formation of an oxide barrier. RECENT investigations1-3 have demonstrated that environmental factors are frequently responsible for delayed failure in martensitic high-strength steels. The environment causes crack initiation from preexisting cracklike flaws at unexpectedly low stresses, often far below the uniaxial yield strength. The crack then propagates at a rate controlled by the environment and the crack-tip stress-field intensity; during this period crack propagation will cease if either the stress or the aggressive environment is removed, and this may be considered a definition of subcritical crack growth. Fracture occurs when the crack has grown to a critical length, such that the fracture toughness of the steel is exceeded. Several liquids are known to cause delayed failure, but water and water vapor are the most severely damaging of the environments investigated to date.'-3 It has been suggested that hydrogen produced during corrosion reactions may be the actual embrittling agent2 The severe embrittlement caused by water vapor inevitably raises the question as to whether other gas environments are damaging. Previous work4'5 has demonstrated that high-pressure hydrogen (2250 psi or greater) is an effective embrittling medium, as measured by such criteria as reduction in area, notch tensile strength, and time to failure. Moreover, the embrittlement was eliminated by the addition of small quantities of air or oxygen to the high-pressure hydrogen, but it was not influenced by additions of nitrogen or argon.4 In the present study attention is focused upon gas environments at atmospheric pressure, in particular hydrogen and oxygen. The influence of these constituents upon the initiation and growth of a crack from a pre-existing cracklike flaw in a high-strength steel was determined directly by the electric-potential method of monitoring crack growth. MATERIALS AND PROCEDURE The H-11 steel was supplied by the U.S. Naval Research Laboratory as 3 by 12 by 0.065 in. sheet specimens with 1-in. transverse center slots. The slot tips were finished by Elox machining to a radius of 0.001 in. or less. The yield strength was 230,000 psi; chemical composition and heat-treatment schedule are given in Table I. The experimental apparatus, precracking procedure, and incremental loading sequence have been described previously.3 In the only significant change, precracking was sometimes carried out by a hydrogen embrittlement and baking technique. All experiments were conducted at room temperature; the hydrogen gas environment was purified by a commercial palladium purifier. Crack initiation and propagation were followed by the electric-potential method,'-3 using a calibration that has been recently dicussed. Stress-field intensities at the crack tip were computed by the conventional Irwin frmula; the plasticity correction was negligible and therefore not included. EXPERIMENTAL RESULTS A striking brittleness associated with molecular hydrogen at atmospheric pressure is demonstrated by the data in Table I1 and Fig. 1, which indicate that subcritical cracks in hydrogen initiate at lower stress-field intensities and propagate more rapidly than in fully humid argon. Ki and Kf are the crack-tip stress-field intensities at initiation and termination of subcritical crack growth, respectively. Termination, of course, corresponds to fracture. In both hydrogen and humid argon Ki is less than Kf

Jan 1, 1967