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By C. G. Dunn, F. W. Daniels, M. J. Bolton

J. P. Nielsen—The data that Dr. Dunn and his associates have been obtaining are welcome checks on the theoretical aspects of grain boundary energies. With reference to the comments on the validity of the inter-facial tension concept, I should like to say that I have been having considerable difficulty myself in untangling the confusion between "interfacial tension" and "interfacial energy." The confusion however resolved itself for me when I learned that "interfacial tension" for solids, or at least crystals, is used with two distinct meanings. One meaning12. ' "hat is implied for "interfacial tension" is the work necessary to produce or increase the surface of the solid in question by a unit area, temperature and pressure held constant. This comes from the somewhat surprising but simple fact that surface tension and surface energy in the case of liquids are identical quantities. This synonymity has carried over to crystals, where it should be kept in mind that the tension connotation has meaning only by analogy with liquids and that no real stresses are being referred to. The only reason that there is apparently a set of stresses acting at a grain boundary junction is the optical illusion that liquid interfaces are being observed for the grain boundaries. The other meaning3,14,15 for the term has to do with actual states of stress in the surface of a material as would be measured by a system of forces applied to the surface to equate the surface stress state with that in the body of the crystal. J. B. Hess—Since the authors have proposed that the surface tension concept should be discarded in order to deal with the general case of interface energies in solids, their attention should be called to a recent paper by Gurney." The latter has shown, from a thermodynamic argument, that surface tension is a valid concept in solids. In fact, Gurney has rejected the idea that surface tensions are merely mathematical fiction and, instead, has given an interpretation of their physical reality. Finally, he has concluded that, as long as appreciable atomic migration takes place, specific surface energies and surface tensions must be equivalent in solids as well as in liquids. Accordingly, the authors' eqs 2 and 3 can be written with equal accuracy in terms of interface tensions, simply by replacing each E,! with its numerically and dimension-ally equivalent Interfacial tension ?,,. If the interfacial tensions (or specific energies) are independent of interfacial orientation, as is true in liquids, then the general equations that define the geometry at equilibrium of the common junction of three interfaces reduce to eq 1. However, neglect of the effect of boundary orientation is never completely justified theoretically' in dealing with crystalline solids; hence, use of eq 1 in such cases must always be considered as an approximation, though clearly a reasonable one for many purposes.1,2 At the same time, one should not be surprised that the assumption that interface tensions in solids are free from orientation

Jan 1, 1951

By Walter Miller

A surprising development during the year was the continued high demand for petroleum products and the high level of crude oil charges to the stills. Generally speaking, the petroleum industry during 1946 succeeded in producing a greater quantity of crude oil and petroleum products than in any of the war years, and considerably exceeded the production of any prewar year. Daily crude still charges averaged between 4,700,000 and 4,800,000 bbl per day for 1946 and about equaled the 1945 record, although most pessimistic 1945 guesses, predictions, or fears ranged, for 1946, from only 3,900,000 to 4,400,000 bbl per day. In spite of the high crude charge to the stills, total stock of crude oil was higher at the end of the year by some 7,000,000 to 8,000,000 bbl. A similar story holds regarding stocks of the principal refined products. Gasoline on hand at year's end was two to three million barrels lower, and was considered low for the expected demand during the coming high consuming season, while stocks of light and heavy fuel oils, although some higher than a year previous, hardly reached a comfortable level. The trend in gasoline of high octane number, which has resulted in a marked rise in. the antiknock quality by the end of 1945 (five or six numbers higher than during the war and one number higher than the prewar quality of civilian motor gasoline of 1941) was abruptly arrested by an acute shortage of metallic lead for manufacturing tetraethyl lead. Government action in the form of lead price control, rationing, and the setting of octane ceiling specifications to conform to the reduced ethyl lead compound available were enforced until abrogated at the end of the year. It will be interesting to see how rapidly the octane levels will resume the rising trend, as more metallic lead slowly becomes available for making greater quantities of ethyl lead compound. As octane quality rises over present levels the disadvantage of the refiner without catalytic cracking equipment will increase. Gasoline made by catalytic cracking, without addition of ethyl lead, approximates 77 octane by the official ASTM Motor Method test, compared with 66 for gasoline generally similar in volatility made by the older thermal cracking operation. Catalytic gasoline shows an even greater improvement in the Research octane number, the comparison being roughly 86 Research versus 71 Research octane for the thermal cracked product. This higher differential in Research octane is credited by many, although not universally so, as contributing to better road performance, and consequently resulting in a more desirable product marketwise. In some parts of the country a premium of roughly one fourth cent per gallon is now being paid for gasoline with higher Research octane made possible by including catalytic cracked product in the final blends. If such price differentials in favor of catalytic gaso-

Jan 1, 1947

By R. J. Mellen

In a chloridizing roasting and cyanida-tion plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42-in. rolls, the latter in closed circuit with two Tyler type 400 screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼-in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the rushed Ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1947

By John K. Hammes

INTRODUCTION From the point of view of the consumer, the cost of mineral commodities might be viewed as the total price industry pays for mine output. Similarly, the mining company engaged in the ongoing business of finding, developing and operating new mines is concerned very much today with the distribution of all the revenue paid for mine output and not just with its direct or controllable cash costs. In other words, what will the labor force, the supplier of energy, explosives, and reagents, and the supplier of government services, management, capital, and the ore reserves receive for their contribution to production? A significant amount might be categorized as what economic theory classifies as economic rent. Mining companies are finding it increasingly difficult to negotiate concession or work contracts to develop ore reserves in lesser developed countries, just as they find themselves accused of taking excess profits in their own countries. The principal difficulty in reaching mutually satisfactory agreements in many cases is the difficulty in establishing an acceptable distribution of income, in particular the distribution of economic rent. (See Chapter 7). Further, it is logical that the decisions and policies of the interested parties be they government or industry should be and will be influenced by their competitive position relative to determining the distribution of economic rent. The raising of finance for a mining project today may be inhibited by too biased an extraction of economic rent by entities other than the risk takers, the mining companies. ECONOMIC RENT A British economist (Ricardo D. McGraw-Hill, 1955) first proposed the concept of economic rent (See Drechsler, H.D. earlier in this Chapter.) Ricardo pointed out that no rent would be paid until land became scarce. Ricardo also stressed the distinction that the price of farm produce was not higher because of the rent paid on farm land but that the rent or price of more fertile farm land was higher because the price of produce is higher. This led to the concept that economic rent is not a cost and that it is therefore not price determining, but is itself price determined. Economic rent may be viewed as a cost if cost is equated with the concept of payment or distribution of income to any factor used in production. However, if it is a criterion that a payment must be made to the supplier of the factor of production, then economic rent is not truly a cost because it is not necessarily distributed or "paid" to the supplier. Consider the commodity A which occurs in three grades: Al, A2, and A3. Any grade may be used as a substitute for another grade in all of the industrial uses of A. However, in industries X and Y the different grades have different productivities. Table 1 shows the relative productivities of each grade in industries X and Y, and in all other industries [~]. [ ]

Jan 1, 1985

By C. W. Donnelly, Louise B. Freeman, Coleman D. Hunter

It is with pride that the authors of this paper report. that during 1944 the production of petroleum in Kentucky passed its all-time peak, 9,496,985 bbl. being contributed. The delivery of natural gas, especially from eastern Kentucky, was maintained and possibly exceeded the highest deliveries of the past. To raise the production of oil and gas was not easy with inferior and worn-out equipment and the shortage of skilled labor. Reminiscent of the early days of the industry in the state, particularly in the hills of Elliott County, where mud roads often have gradients of 45º, was the replacement of worn-out trucks and tractors by oxen. Of the total 9,496,985 bbl., 17 per cent was contributed by the part of Kentucky that is east of the Cincinnati arch, as compared with the total of 7,010,776 bbl. contributed in 1943, of which 25 per cent was contributed by eastern Kentucky. This discrepancy is due to the discovery of more new fields in western Kentucky rather than to a decrease in production in the eastern fields, where the average has been maintained by reconditioning and secondary recovery of old fields and further development of new pools reported for the first time in 1943. In all, 1103 wells were known to have been drilled in the state during 1944, of which 659 were drilled in western Kentucky and 444 in eastern Kentucky. The exact number of wells drilled for any year is never known, as drilling permits are not required except 'when it is planned to drill through a workable coal seam. Of the total number of wells drilled, 459 were oil wells, 229 gas wells and 415 dry holes. The monthly production was as follows: January........................... 659,496 February.......................... 658,606 March...................... ...... 711,619 April............................. 649,926 May............................... 754,892 June.............................. 730,369 July.............................. 761,762 August............................ 878,053 September...................... 863,150 October........................... 968,732 November........................ 961,999 December........................ . 898,636 Western Kentucky In western Kentucky 352 wells were completed as oil wells with a total initial production of 47,922 bbl. of oil, the average being 136 bbl. per well. Two wells produced gas, the total initial production being 1,594,000 cu, ft. One hundred and eighteon wildcat wells were drilled, of which 107 were dry holes and 11 were oil wells. Three of the latter proved to be the discovery wells of commercial pools. Twelve new pools were discovered during 1944, none of major importance, as indicated in Table 2. The Utley pool, due to a stratigraphic trap on the flank of a closed structure, found oil in the Mcclosky at 2565 ft. in the discovery well, with initial productioil of 1000 bbl. The average daily production is now 227 bbl. from seven wells. The East Poole field, with the discovery well in the Tar Springs sandstone at 1811

Jan 1, 1945

By J. H. Starkey, P. A. Scinto

"In Canada and other countries, legislation dealing with accurate factual disclosure for mining projects is being expanded to include more metallurgical design information. Grinding characteristics, metallurgical recovery of contained metal and environmental hazards associated with tailings disposal will be part of the Canadian package. This paper will examine only that portion of the plant design standard practice that relates to grinding – in SAG mills and ball mills that grind ore from as received crushed material to final ground product that is fed to downstream plant recovery processes. Since grinding plants represent a significant portion of the overall concentrator cost, accountability in grinding design should no longer be left to chance, opinion, or hidden proprietary information. This is the 21st century. It is reasonable to expect that a factual basis will be needed so that a Qualified Person (QP) reviewing the grinding mill design can understand it. That will also involve how many and what quality of samples are taken to define the characteristics of the deposit, the accuracy of the tests used to measure ore hardness and the method by which the dimensions and power of the mills are determined. If these criteria are not met, the QP may have no choice but to call attention to the lack of convincing information regarding the grinding mill design for the project. At the same time clients are expected to demand getting the correct design information at a cheaper price, results in a more timely way, and accurate design to the point where the risk of tonnage shortfalls are reduced or eliminated.Tables will be presented to show how this can be done to current standards using existing technologies. At the same time it is acknowledged that current standards are ill-defined and that to achieve consensus among those involved in SAG mill design work, it may require that a consortium funded project be set up to establish what tests and methods for grinding design are suitable to meet the new standards for metallurgical disclosure. By running parallel design methods on several major projects over a period of 5 years, the process of how to design grinding circuits will be better understood because the projects will start at the design stage, include procurement of the mills and will end with a benchmark test on the plant.Selecting the samples and the best schedule for doing test work on core samples submitted and mill design, will be discussed in this paper. The issue of completing the grinding design in time for early procurement decisions is on the critical path for most major projects. Discussions between mining engineers, geologists and process engineers are needed early in the development cycle, at the scoping study level (SS), preliminary engineering assessment (PEA), prefeasibility study (PFS), and at the feasibility study stage (FS), to ensure that proper planning for design sample selection is done. Using adequate sampling and testing methods in a timely way allows early procurement of the grinding mills and minimizes project risk."

Jan 1, 2012

By J. W. Jang, I. Iwasaki

Introduction Grinding ball wear in wet grinding is due to a complex combination of abrasion, corrosion, impact and their interactions. The ball charge is continually falling from a height greater than half of the ball mill's diameter, and contact is made with ore particles or other balls. Abrasive wear occurs through penetration and plowing of media materials from the ball surface by hard ore particles (Gangopadhyay and Moore, 1985). Electrochemical interactions between the abraded and unabraded areas of the ball surfaces, as well as between the grinding media and the minerals, occur in grinding mill pulps (Adam et al., 1986). The wear depends on the media composition, the hardness and the corrosive/abrasive characteristics of the slurry. The marked-ball-wear test method was used to evaluate the performance of grinding media under nitrogen and oxygen atmospheres. Also, the wear mode was studied by observing the grinding media surface under a scanning electron microscope. In general, grinding media of high hardness and high chromium contents were found to have better corrosive wear resistance than those of low hardness and low chromium contents (Jang, Iwasaki and Moore, 1988). The purpose of this investigation was to determine the size distribution and the shape of wear debris in a study of the wear rates and wear behaviors of grinding balls. Experimental Materials The grinding ball alloys used in this investigation were: •martensitic stainless steel (MSS; 0.96% C and 16.2% Cr), •high-carbon, low-alloy steel (HCLA; 0.89% C and 0.53% Cr), and •mild steel (MS; 0.22% C and 0.09% Cr). The mineral used in this study was synthetic corundum, because it was on hand and contained no magnetic impurities. Debris A 203- mm (8-in.) diam porcelain mill was used for grinding the corundum. The mill charge consisted of: •1200 g of 12-grit synthetic corundum, •770 mL of 0.05 mole/L Na2SO4 (61% solids), and •125 steel balls of nominally 25-mm (1-in.) diam. To minimize the debris corrosion, nitrogen was introduced into the mill. Grinding with MSS, HCLA-steel or MS balls was conducted for 2 hr at 51 rpm. The grinding time using the MS ball was varied from 0.5 hr to 2 hr in an attempt to study the debris amount and shape as a function of time. The debris was collected with a hand magnet from the ground slurry and separated from the corundum particles using a Davis magnetic tube. After the separation, the debris particles were immediately washed with acetone and then filtered and dried. Debris size analysis Buckbee-Mears 3-in.-diam, micro-mesh sieves and an electromagnetic shaker were used for determining the debris size distribution. In sieving, isopropyl alcohol was used. The debris fraction collected from each sieve was washed with acetone, dried and then weighed on an analytical balance. [ ]

Jan 1, 1993

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION This method is a variation of cut-and-fill mining where either the wall rock or the vein is too unconsolidated to permit safe mining by conventional cut-and-fill methods. Basic development consists of driving a 2.4 x 2.7-m (8 x 9-ft) haulage heading laterally in the footwall side of the vein. Crosscuts are driven over to the vein at 91-m (300-ft) intervals. Then the vein, at each inter¬section, is mined out to a distance of about 9 m (30 ft ) above the track level. After the installation of an ore chute, scrams or intermediate drifts are driven in both directions along the vein. Mining then starts imme¬diately below the level bounding the upper limit of the stope. A cut varying between 3 and 4.2 m (10 and 14 ft) in depth and mined for a distance of 45.7 m (150 ft )along the strike is taken from the orepass in each direction (Fig. 1). In this mined-out void, the floor of the cut is equipped with 203 x 203-mm (8 x 8-in.) caps hitched into the wall with steel saddle hitches installed on 1.5-m (5-ft) centers. Double wood stringers, 152.4 x 152.4 mm (6 x 6 in.), are then installed on 1.2-m (4-ft) centers longitudinally down the length of the stope. Reinforc¬ing wire mesh is also placed, overlying each mat of timber. Directly above the stringers, stulls are installed from hanging wall to footwall to hold the timber in place during the sand-filling operation. Then a cement¬sand mixture in a ratio of about 1:10 weight is poured to a depth of about 2.4 m (8 ft). The entire cycle is then repeated down to the scram drift. Another variation of timbering is illustrated in Fig. 2. There is a vein width of approximately 5.5 m (18 ft) with the same unconsolidated ground conditions. The primary difference between handling this situation and the previously described undercut-and-fill method is that standard timbered sets are installed in each excavated stoping void rather than stulls. Once again the basic method is to work down the ore block from an upper timbered drift to a lower timbered drift. Cuts are taken from a timbered raise with a chute on the stope side. Each piece of ground excavated is essentially two sets high, 5.5 m (18 ft), by two sets wide, 5.5 m (18 ft), by the length of the cut along the strike of the vein. The timbered sets are 2.4 x 2.4 x 2.4 m (8 x 8 x 8 ft). The net result is, therefore, two sets of timber on a top floor separated by doubled-up 76-mm (3-in.) lagging from

Jan 1, 1982

By David M. Parkes, Mike Fisher

In 1947, coal production in Western Canada was 10 Mt (11 million st) which predominantly came from 120 underground mines. By 1968 production was down to 1 Mt (1.1 million st). Production in 1981 was back up to 29 Mt (32 million st), but only 1.5 Mt (1.7 million st) of the total was from underground mines. In fact, because of competition from open-pit mines, only two underground mines were operating in 1981-the McIntyre and B.C. Coal mines. While several other underground mines are being planned, this article will deal with only the two in production. Geology The geology in Western Canada is frequently complicated; hence only more regular deposits are mined. Coal is generally thick, often steeply pitched and usually soft or well cleated. The 5000-km (3,100-mile) coal belt that runs from the Yukon to Mexico is well developed in Western Canada. It is divided into two distinct regions-the Rocky Mountain Lower Cretaceous coals and the mostly uniform level Upper Cretaceous coals of the plains area to the east. Isolated pockets of coal occur at Vancouver Island and Hat Creek. Hat Creek's massive 300-m (984-ft) thick geologically recent lignite deposit is 250 km (155 miles) inland. Geography The infrastructure of Western Canada is shown in Fig. 1 along with the locations of the two mines. Both rely on exports of premium low volatile metallurgical coal mostly to the Pacific Rim. The rail distance from both mines to port is 1100 km (684 miles)-accounting for a significant portion of the sales cost of the coal. Mining Methods Two main mining methods are used-room and pillar and hydraulic. Longwall has been attempted and is being considered again. In the past, a simple hand method used gravity to help mine most of the coal on steep pitches. Levels were driven in the seam, rising slightly to assist in draining. On the advance, mining would start at the upper level, usually from adits driven in the side of the mountain. Rooms or angles were driven up 100 m (328 ft) at a suitable angle for the coal to flow down. On retreating out of the level, the pillars would be taken, starting from the highest point, using batteries to stop coal or rock flushing down the room. Batteries consisted of three to six stout wooden posts set at an angle to resist material flow. Coal loading was directly from chutes at the bottom of raises into 1- to 2-t mine cars hauled by diesel or battery locomotives. Often output was restricted by the supply of empty cars. Mechanization began in the 1960s but was difficult in these conditions. Mines such as Canmore and Coleman continued operation into the 1970s but closed because steep seams, high methane emissions, and faulting made mining difficult. McIntyre Mines The McIntyre mines are located 360 km (224 miles) west of Edmonton, Alberta, in the foothills of the

Jan 4, 1983

By H. A. Doerner, F. D. DeVaney, J.B. Clemmer

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By R. J. Wasilewski

THERMAL expansion of rhenium data have been reported by Agte et al.,&apos; and Medoff and cadoff,&apos; respectively, while the linear expansion coefficient was determined by Sims et Al.3 Denoting expansion coefficients parallel and at right angle to c axis by all and Al respectively, their weighted mean should be equal to dilatometric linear coefficient of a random polycrystalline specimen. The values reported for these coefficients are tabulated below: Since these values are in poor agreement, and because of anomalous expansion behavior observed in titanium 4 and chromium, a rede termination of the axial expansion was carried out. A Unicam 19-cm-diam high-temperature camera was used, with unfiltered CuK radiation. Camera calibration was carried out using at Pt wire standarde at temperatures up to 1100oC, the variation between indicated and true temperature being within the experimental error (estimated at <3oC) above 750o C. Rhenium Johnson Matthey spectrographic standard) was vacuum annealed at 1000 oC prior to X-ray diffraction experiments. Lattice parameters were obtained by least-squares solution using the reflec- tions (1232), (1015), (2024), and (3030) for all the patterns, and the additional reflection (2133) for higher temperature (400°C and above) patterns. The relative expansion data obtained are shown in Fig. 1, lengths of the vertical bars indicating the spread in the parameter values obtained by using various combinations of three spacings from the four or five measured. The data calculated from the equation of Ref. 2 are included for comparison purposes. Several specimens were used to obtain the above data, neither of them showing a significant parameter change after completion of the high temperature runs. Some parameter data and the corresponding expansion coefficients between 25 oC and the measurement temperature are tabulated below. The comparison of the present data with those previously reported shows reasonable agreement with the results of Medoff and cadoff2 within the temperature range covered by them (up to 1000 oC). The a values are appreciably higher throughout most of the investigated temperature range. The basal expansion is—in agreement with their findings—appreciably higher than that in the direction of hexagonal axis. The expansion coefficient, al, however, decreases rather more slowly with temperature than previously reported. The expansion along the c axis up to nearly 1000°C appears in excellent agreement. However, at higher temperatures there is a marked deviation towards higher expansion, reproducible between different specimens. Furthermore, there seems to be some irregularity in the variation of all with temperature, the apparent local minimum and maximum values at 510o and 1024o C, Table II, being well outside the experimental error limits. It must be stressed that the apparent discontinuity in the c parameter expansion is still insufficient to account for the expansion value reported by Agte et al.&apos; since a total expansion of close to 2 pct would

Jan 1, 1962

By Ralph M. Hunter

Until recent years, maintenance of surface coal handling facilities was a relatively simple task. Equipment consisted principally of conveyors, screens and crushers of comparatively simple construction. Electrical equipment consisted of motors and starters, with little, if any, interlocking. In most cases failure of any component affected only the grades of coal loaded so that even a prolonged breakdown did not affect mine operation materially. However, with the introduction of modern cleaning plants, the picture has changed. The equipment is much more complex and therefore more difficult to maintain. In many cases failure of any component of the plant prevents loading of coal. Even if provision is made to bypass, the quality of coal loaded is lowered, affecting markets. In the case of plants designed to permit full seam mining, marketable coal cannot be loaded without operation of the cleaning plant. Under these conditions it is apparent that considerable thought and effort toward better maintenance are justified. In order to assure continuous, efficient and economical operation, maintenance should receive consideration, not only after the plant is in operation, but while it is being designed. It is hardly necessary to say that the designer should strive for sturdi-ness and wearing qualities. It is sometimes not realized, however, that even the best parts will sooner or later require replacement or repair. The designer should, there- fore, give consideration to accessibility. Anyone who has had to do with cleaning plant maintenance is familiar with installations where several hours are required to get at a part that can be repaired or replaced in a few minutes, once it is accessible. As in most cases, experience is the best guide. The men who operate and maintain this equipment know the problems encountered. Whenever available, the opinions of these men should be sought and heeded. As an example, we might consider the refuse elevator on a Chance Cone. Our company is now operating three of these cones, the first of which was installed in 1941, and the latest in 1946. Exact records are not available, but well over 50 pct of delays charged to the Chance Cones have been caused by the refuse elevator. This elevator receives refuse from the cone and carries it to a point above the water level of the cone, discharging onto a dewatering and desanding screen. Since the elevator operates under water and necessarily has close clearances in the lower boot, it tends to jam on large pieces of refuse or tramp iron. The chain, buckets and boot are also subjected to the action of water, sand and refuse, which accelerate normal wear. In case of jamming or chain-breakage it is necessary to drain the elevator and the cone so that even a minor fault in the elevator causes a considerable loss of time for the cone. In our new plant, which will contain two Chance Cones, the refuse elevator has been eliminated. This was accomplished by raising the cone so that the refuse gates could discharge directly onto the dewater-

Jan 1, 1949

By J. B. Clemmer, F. D. DeVaney, H. A. Doerner

Production of magnesium metal in the United States during the past decade has increased from less than 600,000 lb. in 1928 to more than 4,800,000 lb. in 1938.1 The growing industry has stimulated interest in methods of production and in sources of raw material. All of the present domestic production is from brines. However, the availability of large tonnages of magnesite in the state of Washington, together with the possibility of cheap power, have created interest in the utilization of magnesite for the production of magnesium metal as well as for other purposes. Magnesite, the carbonate of magnesium (MgCO3), usually is associated with limestone (CaC03) and dolomite (MgCO3. CaC03). It occurs in crystalline and microcrystalline (sometimes termed amorphous) forms.2 The more common crystalline mineral is found in various shades of white, gray, yellow and brown, the colors being due to impurities, chiefly iron. Magnesite has a hardness of 3.5 to 4.5 and a specific gravity of 3.0 to 3.1. It forms rhombohedra1 crystals in the hexagonal system, having a vitreous, pearly luster. The quarries and mines of the Northwest Magnesite Co. near Chewelah, Wash., are large producers of magnesite for refractories. The deposits are extensive, some veins being hundreds of feet wide. The boundaries of the veins are indefinite, and the magnesite is interfingered with dolomite. Calcite, talc, serpentine, quartz and shale are disseminated through the deposits. The quality of the ore cannot be determined by appearance or by simple tests; chemical analysis is required. These features make it difficult to determine the boundaries or total quantity of workable ore.3 Only the higher-grade ores are mined, and a considerable proportion of the magnesite content is discarded by hand-sorting and log-washing methods. After calcining, the sorted product meets the specifications for usual refractory material, as it contains approximately 81 per cent MgO, 7 per cent SiO2, 6 per cent CaO and 6 per cent Fe203 + A1203. A sample of the reject material had the following analysis after calcining: 65.4 per cent MgO, 6.3 CaO, 21.1 SiO2 and 6.5 R2O3. The percentage reported as R2O3 includes compounds whose chlorides are acid-soluble but give precipitates in ammoniacal solution that are ignited and weighed. The more common constituents are Fe2O3, A1203 and TiO2. The sorted magnesite product is not the best grade. even for refractory purposes, and leaves much to be desired for a reduction process. The rejected material contains a large proportion of magnesite. It is apparent that an effective concentration method would yield a higher recovery and a better product. In general, the refractory grade is not suitably pure for other uses. Recent developments in the production of

Jan 1, 1943

By N. J. Grant, Hans Brunner

ThE materials used in this investigation were furnished by the Aluminum Co. of America and consisted of high-purity alurhinum (99.995 pct) and two aluminum-magnesium alloys containing approximately 2 and 5 pct Mg. The compositions of these alloys are listed in Table I. The materials were received as hot rolled 3/8-in. diam rods and were made from the same heats as the materials which were previously used in this laboratory.&apos;j3 The rods were cold swaged to 1/4-in. diam by 1 in. long, and then had two parallel flats machined over the gage length 0.085 in. apart. The test bars were annealed for 10 min at 700° F, and strained 1 pct in tension at room temperature. They were electropolished (80 pct acetic acid; 20 pct perchloric acid, 70 pct concentration) at a temperature of 60°F at approximately 20 v. A grid, transverse or square, spaced 0.2 mm apart, was impressed by rolling a fine threaded screw (128 threads per in.) across the flat surfaces. The depth of the grooves was less than 5 p. For final annealing and grain size control the specimens were enclosed in aluminum foil and sealed in Vycor tubes. Annealing conditions varied from 1 hr at 700° F to 5 days at 1050° F, depending on the alloy content, to give for most of the specimens a final grain size of about 1 mm. This treatment was followed by electropol-ishing as described above. For the test series presented in Figs. 4 and 5 round specimens were used instead of flat ones, and fine machining marks served as marker lines instead of the impressed grid lines. Electropolishing was omitted. The major advantage is that the spacing of the machining marks is only about 1 u as compared to the 200 u of the impressed grid lines, permitting higher precision in strain measurements. Unfortunately the strain annealing necessary for grain size control also resulted in some preferred orientation of the grains. The preferred orientation of three arbitrary specimens was measured by X-ray diffraction methods and the averaged results are shown in Fig. 1. The possible effects of the preferred orientation on the test results will be discussed later. The testing temperatures were 410°, 515", 725", and 940° F, with one exception in the case of the 5.1 pct Mg alloy. Since the solubility limit of this alloy is 520°F, a test temperature of 545°F was cho-se sen instead of 515°F. The selection of the test temperature was based on two considerations: 1) at 900°F and above, extensive grain boundary migration made it more difficult to measure grain boundary sliding; 2) at 400°F and below, there was so little grain boundary sliding that it was again difficult to make precise measurements. The stress was chosen to give a creep rate of approximately 2.7 pct per hr in connection with Figs. 2

Jan 1, 1961

By R. G. Vardiman, M. R. Achter

THERE has been interest recently in the conditions under which dislocation pipe diffusion may be observed. In order to extend the measurements to temperatures below Tm/2, where pipe diffusion becomes more important, a chemical sectioning technique was developed for silver single crystals which makes possible the removal of layers of 100A or less. The procedure is briefly as follows. Silver single crystals of 99.995 pct purity were cut in the form of discs t by 4 in., electropolished, and electropolated with Ag110 to about 1000A thickness. The thickness was determined by calibrating the activity with the weight of heavier platings, up to 10,000A. Both diffused and undiffused samples were sectioned in a two-step process. Reaction with a solution of iodine in alcohol formed a thin, even layer of AgI, which was then dissolved in a KCN solution. The reaction rate of the iodine solution varied slightly over a period of time, so that the thickness of the stripped layers, while quite even and regular for a given specimen, ranged from 65 to l00W for different specimens. This variation was checked by using a thickly plated standard with each run. Layer thickness was determined from undiffused samples by the point where the layer activity drops to half the maximum. This maximum intensity also provides, with appropriate corrections, the initial concentration per layer co. Because of the small penetration distances, it is necessary to use the exact solution of Fick&apos;s law for a source of finite thickness.&apos; For a plating of thick- where c is the concentration at depth x. flj the error function, D the apparent diffusion coefficient, and t the time. The diffusion coefficient was obtained from the concentration-penetration data by extrapolating to x = 0 where c/co = erf h/2(Dt)1/2 Enough points were taken to impart confidence in the accuracy of the intersection, generally about ten layers. Four temperatures were used as in Table I. For best accuracy, t and h were chosen to give values of the parameter (Dt)1/2/h between 1.0 and 1.7. A typical concentration-penetration curve is shown in Fig. 1. The diffusion coefficients calculated from Eq. [I] are plotted in Fig. 2. For comparison, the high-temperature data of Tomizuka and sonder 2 which go down to 630°C are plotted as the solid line 2(a) from their expression D = 0.40 exp (-44,090/RT). Also for comparison, the lowest temperature point of Hoffman and Tumbull3 is plotted. It is apparent that our data diverge appreciably from the high-temperature extrapolation by 373°C. A better straight line fit for the low-temperature portion is obtained by subtracting the extrapolated volume diffusion coefficient from the apparent. or measured coefficient, and plotting the difference, curve (h) of

Jan 1, 1970

By James A. Finch

The Gaudin lecture is devoted to fundamentals. Given my involvement with column flotation over the last 20 years, this is my topic. I have tried to identify spinoff basic research and provide a personal list of favorites. Two are selected for discussion here. They are single bubble rise and bubble surface area flux. The rationale for such work, as well as some of the findings, are reviewed. Support for fundamental work seems as volatile as the economy (and the link is probably not coincidental). I hope some utility to the minerals industry is evident in the text. Column flotation in its current form originated in the early 1960s. The method was invented by Boutin, patented by Boutin and Tremblay and pursued doggedly by Wheeler. Their aim was to overcome the problem of mechanical entrainment of fine particles in the water carried by the bubbles, which they were convinced contributed to lowering flotation concentrate grades (Wheeler, 1988; Finch and Dobby, 1990). The solution was a combination of columnar geometry, the addition of wash water into a deep froth and the maintenance of an excess of water to the tailings, as compared to the feed (a so-called positive bias). Success at Les Mines Gaspe, Quebec, on cleaning Mo in the early 1980s signaled the start of industrial acceptance, making the column one of the flotation machines of choice today. Over the next 20 years, column applications mushroomed to cover virtually all mineral commodities, as well as nonmineral applications such as soil decontamination, deoiling of water and deinking of recycled paper. In the wake, research groups are continuing to probe the fundamentals, conferences and workshops have been transferring the technology and specialist companies have emerged to service the demand. Innovations were triggered as problems arose (for instance, in gas injection systems) and novel flotation cells, much in vogue in the 1990s, had the door opened. Mechanical cell manufacturers have not been idle - there is nothing like competition to provoke a fresh look at an established business. Flotation-cell technology has probably not seen so much activity since the patent rush in the early days of flotation, which occurred around the turn of the century. For an industry often considered reluctant to risk innovation, this is not a bad record. A good idea will find a place, but patience (not always an innovators strong suit) and careful preparation are required. Two books (Finch and Dobby, 1990; Rubenstein, 1996), three symposia (Sastry, 1988; Agar and Huls, 1991; Gomez and Finch, 1996), and a series of reviews (Dobby and Finch, 1991; Finch and Dobby, 1991; Finch, 1995; Finch et al., 1995) document the period. Like many good ideas, column flotation has been overexposed, and there have been disappointments. In a recent example, columns performing scavenging duty were replaced by mechanical cells (Chong et al., 1998). The further applications stray from the original - to

Jan 1, 1999

By P. R. Sperry

Aluminum alloy 3S was examined to determine the relationship between the applicable phase diagram and the microstructures produced under conditions tending toward non-equilibrium as well as equilibrium. The only phases present in the alloy were aluminum, (Mn,Fe)A16, aAI(Mn,Fe)Si, and silicon. It was found that the alloy behaved essentially as though it belonged to the Al-Mn-Si system. The effects of various solidification rates and of various homogenization treatments were studied. Particular attention was given to a nonuniform precipitate distribution which reflected the solute distribution in the cast aluminum solid solution. Some possible explanations for the solute distribution based on characteristics of the Al-Mn-Si or AI-Mn systems were discussed. ALUMINUM alloy 3s (aluminum plus 1.25 pct Mn) is one of the oldest and the most widely used of the wrought aluminum alloys. In spite of this fact, there are some aspects of the metallography of 3s which have not been revealed or which have not been brought to the attention of the metallurgist working with the alloy. The purpose of this investigation was to demonstrate in general how some of the fundamentals of equilibrium phase diagrams may be applied to the study of an alloy and to show specifically the origin of certain metal-lographic features of aluminum alloy 3s. It is hoped that this information can serve as a basis for solving specific problems relating microstructure to such things as mechanical properties, recrystallization, grain growth, and surface finish. Literature Survey Considering that alloy 3s contains only one alloying addition, manganese, it seems reasonable that it would have characteristics pertaining to the A1-Mn system. However, those characteristics might well be altered by the impurities in the commercial alloy, especially iron and silicon. Therefore, not only the simple binary but also more complex systems had to be reviewed. One characteristic of the A1-Mn system worthy of mention is its sluggishness, which leads to conflicting solubility data1,&apos;2 and makes it relatively easy to supersaturate the solid solution and shift the eutectic composition during rapid solidification.3&apos;4 Another characteristic is the strong influence which small amounts of impurities, especially iron, exert in lowering the solid solubility of manganese. Undissolved manganese is ordinarily present in an intermetallic phase, MnAl,, having an orthorhombic crystal structure. A metastable precipitating phase designated as "G" has been reported but small amounts of iron and silicon suppress this phase entirely." Raynor has established that MnA1, can dissolve a considerable amount of iron by a substitution of iron for manganese atoms up to 50 atomic pct.&apos;" Consequently, in the Al-Mn-Fe liquidus diagram, the eutectic valley separating the primary aluminum and (Mn,Fe)Al, fields is fairly close to a constant Mn + Fe content." This is an important consideration in regulating the composition of 3s and also other aluminum alloys with high manganese additions."

Jan 1, 1956

By Everett H. Graff, Ray H. Misener

THE Sol de Oro mine is in the province of Nazca, 487 km. southeast of Lima, and 60 km. inland from the coast at an elevation of 1200 m. above sea level. From Lima the mine may be reached in 8 hr. by automobile; 472 km. from Lima to Nazca over the paved Pan-American Highway and 15 km. from Nazca to the mine over a good graveled road. The trip from Lima can also be made in 2 hr. via the Faucett Air Lines which make two stops weekly at Nazca. The [Compania] Aurifera Nazca, owning and operating the Sol de Oro mine, was organized in Lima in 1930 by a group of prominent Peruvian mining men, including Ing. Carlos Alvarez Calderon, Dr. Alfredo Alvarez Calderón, Ing. Hector Boza, Ing. Jorge Félix Remy, Ing. Eulogio Femandini, Sr. Fernando Wiese, and Sr. Julio East. The nine claims in the Sol de Oro group owned by the company comprise an area of 52 square kilometers. Development of the property was not started until 1932 and only then under adverse conditions, owing to lack of transportation facilities. Mining machinery and other supplies could be taken as far as Ica by boat and railroad, but to transport them over the 180 km. of hot, sandy desert from Ica to the mine was a problem. There was no road in those days and only a few of the most venturesome truck drivers cared to contract the hauling. Another problem the company had to face was that of securing adequate labor. The Sol de Oro is in an agricultural region, and was forced to look to the near-by cotton plantations for men. Experienced men were imported from the "sierras," but they did not stay. The men are accustomed to high altitudes and cold climate and cannot easily acclimate themselves to the hot and arid coastal region. As a consequence the company has been forced to hire the inexperienced local laborers and train them for mining. The company has established three separate camps at the Sol de Oro property. Each is approximately one kilometer from the other and connected by auto road. The upper mine camp consists of living quarters for 300 workmen and their families, a general mine office, carpenter shop and timber-framing shed, compressor house, drill-sharpening and drill-repair shop, first-aid station, time office, school and store. At the central or "empleados" camp there is a 200-ton cyanide plant, general offices, warehouse, powerhouse, shops, laboratory, hospital, school, guest house, an 18-room hotel for the single staffmen, and 12 completely modern five-room cottages for the married staff employees. At the lower camp are living quarters for 200 workmen and their families, two general stores, a school and a small theater. Below this camp, along the river, are the company wells and pumping plant supplying water for mine, mill, and camp uses.

Jan 1, 1945

By M. Coleman, M. E. Reed

The control of feed delivery quality to the flash furnace remains an unsolved problem. Many operations around the world still have limited control of feed quality to their flash furnace. Schenck Process and WorleyParsons continue to work towards the ‘holy grail’ of simple quality feed delivering both radial feed distribution and pulseless feed delivery. Ten years ago they joined forces to offer a system based on then Clyde Process RotoFeed system which could deliver ground level installation, online mixing of materials, pulseless feeding direct to the burner, with good radial distribution. Unfortunately, this proved to be too large a step for the Flash Smelting community and met with resistance from the process owners. Since then the community has developed a range of options some of which have been implemented over this period. Despite these improvements there are still areas of concern including safety issues associated with concentrate leaks and spills, uncontrolled variations in feed rate including massive flooding events, intermittent distribution improvement and set up of the burner continues to be trial and error and hard to repeat the success in each burner campaign. These continue to impact flash furnace performance globally on a daily basis. Following the community reticence to pursue the complete solution a number of interim steps are now available to pave the way to the ultimate feed system. The mixing of material has been installed and is delivering line blended materials on a daily basis. This paper will investigate the developments over the last 10 years and discuss the pros and cons of this incremental step process for overall furnace control. INTRODUCTION Schenck Process and WorleyParsons have continued to interact and observe changes in the flash smelting industry and to consider application of the Clyde RotoFeed now incorporated in the Schenck Process ProStreamTM range for delivery of feed directly to the concentrate or matte burner. Schenck Process pneumatic injection from the Clyde Process range, ProStreamTM, utilises two types of volumetric feeders to accurately control the feed of material into an injection pipeline. These are referred to as the RotoFeed and RotoScrew technologies and the choice of feeder type is dependent on the characteristics of the material to be handled. The RotoFeed is generally used for finer, granular and powdered materials which cannot be accurately handled with a screw such as pulverised coal and copper concentrate, whilst the RotoScrew is particularly suited to coarser lump materials, particularly abrasive of cohesive materials like zinc powder, electronic scrap and metal ashes. Both units can operate at high pressures and many systems work at 3 to 10 bar with the occasional system working at 20 or 30bar for carbon to gasifiers. For the Flash furnace the ProStreamTM units can be used to deliver each of the furnace feed materials into a single line, mixing the materials homogeneously and live to the burner.

Jan 1, 2019

By A. J. Phillips, A. A. Smith, J. S. Smart

THE preparation of a sufficient quantity of high-purity copper for extensive laboratory use in the study of impurity effects has been an integral part of a research program in progress at the Central Research Laboratory of the American Smelting and Refining Co. for the past several years. Initially the degree of purity desired was that necessary to meet two essential requirements for experimental accuracy: (I) no impurity was to be present in the copper in an amount capable of producing a detectable effect on the properties to be studied, (2) even the minor changes in properties caused by the addition of extremely small amounts of individual elements were to be measured without significant interference from contaminants. In the course of the investigation, evidence has accumulated to justify the conclusion that this copper is of exceptional purity; in fact, no impurity can be detected by any chemical method or spectrographic technique known to the authors. The present paper describes the method of production of the copper and presents some of its properties. PREPARATION OF HIGH-PURITY COPPER The elimination of impurities from commercial electrolytic copper is accomplished by a three-step purification process. The first operation consists of an electrolysis through a purified CuS04-H2SO4 electrolyte using bagged anodes, low current .density (5 amp. per sq. ft.) and low temperature (25°C. max.). The electrolyte is prepared from commercial Cu2SO4 and is purified by heating the solution and adding I gram Fe as Fe 2(S04)3 to each 100 grams dissolved copper; sufficient hydrated oxide of copper, prepared from Cu2SO4 and NaOH, is supplied to precipitate Fe(OH)3, which carries down arsenic, antimony, tin, selenium, tellurium and other minor constituents. Silver and remaining traces of lead are removed by addition of Na2S to the acidified filtrate, which is then refiltered after standing hot overnight. The cold solution is made up to 30 grams per liter Cu and 150 grams per liter H2S04 and introduced without addition agents into the electrolytic system, which consists of a double row of hard-rubber battery boxes with the inside webbing and partitions removed, arranged in seven steps of ascending height. Electrolyte is uniformly supplied to the two highest cells by glass siphons operating from a constant-level reservoir, and flows by gravity through the bank of seven cells, entering each at the bottom and leaving at the top. Glass air lifts continuously supply the returning liquid to the constant-level reservoir for recirculation. Each cell contains three 434 by 43a by I-in anodes cast from high-quality electrolytic wire bar copper, various lots of which have been used with equal success. The anodes are pinned to sheet copper hangars shaped to prevent crawl of the electrolyte to the multiple-wired bus bars-which would result in fouling of the contacts and hang inside wool flannel bags stretched over open wood frames. The deposit

Jan 1, 1941