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By C. W. Thompson

Paul D. Suloff (Goodyear Tire and Rubber Co., Inc., Akron, Ohio)—I would like first to comment on problems of the conveyor belt discussed in Mr. Thompson's excellent paper, since that is what we hope we know most about. Twists in relatively wide conveyor belt unavoidably produce a lateral maldistribution of tension, raising tension at belt edges and reducing it at the center. They also produce a lateral collapsing force on the belt at the center of the twist owing to the inherent tendency of all the longitudinal elements of the belt to try to pass through a point at the twist center. Calculation of the twist geometry by the methods shown in Mr. Thompson's paper keeps these extraordinary forces within limits which the belt designer can tolerate. No reduction in belt life due to twisting need be contemplated when this geometry is maintained. There is a minor exception that belts of extreme lateral flexibility will tend to curl laterally at the center of the twist. However, any ordinary fabric construction will perform satisfactorily in this respect. These twists are always made in regions of low tension in the conveyor so that even in the edges of the twist, belt tension does not exceed the average tension found in highly stressed regions of the conveyor. Offsetting these out-of-ordinary belt stresses is the advantage that Mr. Thompson has brought out of getting the return run up out of the dirt where it can be seen. This not only makes it easier to train, but also, in the event that it is not properly trained, frees it of the normal return run edge wear hazard. It is well known that return run edge wear is a prominent cause of belt mortality underground. An interesting aspect of this two-way conveyor is that the belt may be made what is known as a Mobius Strip. A Mobius Strip is obtained by splicing a belt after turning one end of it 180" about its longitudinal axis. In other words, one end is turned upside down before splicing. A belt spliced in this fashion turns itself upside down every time it comes around, but the twist which has been put in the splicing, of course, stays at one location on the conveyor, in this case one of the twist sections at the end. Turning the belt over every revolution might have advantages in some cases. Belts could be made with equal covers and the two sides worn equally and simultaneously. In this case there would be no problem of getting belts on upside down by mistake. However, the two-way conveyor does not have to be a Mobius Strip. It can be twisted in such fashion that the same side is up on both runs. It is simply a question of which way the final 90" twist is made before joining the ends. Another interesting aspect of the two-way conveyor is the problem of operating two-way conveyors in series. Here the sequencing of starting brings up some new problems. It will be recognized, although not always at first glance, that if the starting sequence is planned for one run of the conveyor the reverse will result on the other run. With the two runs carrying bulk material in both directions a reverse sequence on one run would be intolerable. In this situation the only solution appears to be a simultaneous starting of all conveyors in the series. However, with the coal in one direction and intermittent supplies in the other it would be entirely practical to sequence the conveyors for the coal run and accept a reverse sequence on the supply run. The two-way conveyor also lends itself to new driving possibilities. First, it is quite possible to drive at the head end of each run, which of course, means a drive at each end of the two-way conveyor. Driving in this way a given belt can be extended to substantially greater lengths than a conventional conveyor with drive at one end only. In addition to this, under certain conditions the conveyor can be extended to extreme length by driving at one end and at some intermediate point on the most heavily loaded run. As a particular case, a belt carrying coal downgrade and supplies back upgrade could be extended to extreme lengths by driving at the head of the coal run and at an intermediate point of the supply run. Mr. Thompson has been a pioneer in belt conveyor transportation underground and his accomplishment here with the first two-way conveyor of any consequence is another notable addition to the art.

Jan 1, 1954

By E. P. Pfleider, W. D. Lacabanne

Rotary percussive drilling, vibro-drilling, vibro-rotation drilling, rotary pulsating drilling-all names that mean the same thing-are gradually appearing in mining literature. What is rotary percussive drilling? It is a combined system superimposing a rapid percussive impact on a continuous rotary action in an attempt to utilize the most desirable features of both methods. At present single and multiple carriage-mounted rotary percussion units are being used in the European coal mining regions. One well engineered unit is the Hausherr "Albo" rotary percussive drill, utilizing two alternately operating percussion pistons, and another is the multiple-unit, jumbo-mounted Salzgitter machine. Niisse & Grafer also manufacture rotary percussion drills.

Sep 1, 1955

By A. Lee Barrett

Mine power systems are quite different in many respects from those usually found in industrial plants. Wide areas are served, usually by a circuit which is connected continuously throughout the mine. In some mines one circuit may run for several miles. There are usually a small number of branch circuits involved considering the areas covered. In all mines circuits are sectional-ized at approximately 2500 ft by means of switches. All, or at least most branch .circuits, are sectionalized by manual switches. In some mines automatic sectionalizing gear is used to give automatic overload protection at various strategic points in the system. However at the present time this practice is by no means general. Circuit hazards in mine electric systems are much greater than those encountered in industrial installation. Falls of roof or coal form a prominent hazard which may break the wire and may cause ground faults or partial ground faults. Ordinarily electric cables in mines cannot be insulated, therefore a fall can easily cause a short circuit. The humidity in most mines is greater than 95 pct and at many points the floor is wet and water may be dripping from the roof over insulators, increasing the difficulties at that point. All electrical circuits must be contained in the mine entries and space must also be provided for moving equipment and men. Consequently, because of the proximity of the power circuit to the locomotives and other moving equipment a considerable hazard is introduced. Mine haulage track is not always good so that wrecks are not unusual. A wreck frequently does considerable damage to the power circuit. A ground fault frequently results in heavy arcing which may be sustained. If the damage to the circuit is mechanical it may mean that a trip of coal has wrecked and a great deal of coal dust could be thrown into the air as a result of the wreck. The arc resulting from the ground fault could ignite the coal dust and a mine explosion would result. A continued ground fault, even if high in resistance, also may ignite the coal which may be on the bottom of the entry and a mine fire may result. The fire hazard is one of the most serious hazards in any coal mine. Since the mine circuit is completely interconnected, any power interruption resulting. from a fault seriously affects the whole operation. Unless the mine is completely sectionalized the whole mine will be shut down until the fault is located and the proper switches are opened to isolate the fault; even then large areas may still be without power if they are supplied by a single power circuit which may be opened in isolating the fault. Further, mine power is usually supplied from a relatively small number of substations and the fault may be on a stub feed circuit, in which case power cannot be restored until the fault has been cleared. Practice varies widely with respect to placement of substations. Usually they will be spaced from I to 2 miles in 500V opera-tions, and at I mile intervals or less, in 250V operation. Actual mining operations govern substation placements and no general rules can be followed.

Jan 1, 1949

By C. H. Li, R. J. Stokes, T. L. Johnston

Sodium chloride crystals are brittle in the presence of surface microcracks. Water immersion leads to an enhancement of ductility which is retained indefinitely in dry air providing the crystal is dried without leaving a surface precipitate. Storage in a damp atmosphere results in a surface reaction and consequent rembrittlement. Although elimination of surface flaws in ionic solids does not necessarily enhance ductility, there is always a change in the location of the fracture origin. EVER since the original experiments of Joffel there has been controversy over their interpretation and no satisfactory generalized explanation has yet been given. It is generally admitted that when sodium

Jan 1, 1961

By Glenn C. Waterman

THE diamond drill is a very important tool in exploration and development testing and its use is increasing. In almost all cases results of diamond drilling are analyzed on the basis of grade and tons. A proper evaluation of core and sludge assays is important if drilling results are to be acceptable as a basis for geologic and engineering appraisal. The relatively wide variation in assay averages as calculated by various well-known combining methods indicates that the engineering choice of a method may affect the outcome of the drilling in terms of ore and waste. The problem of combining assay results from core and sludge samples has been discussed many times in conference and in the literature.'-' Most writers agree that the field of disagreement in methods is large and that the engineer on the job must consider features unique to his drilling, pick one of several combining methods, and depart from the rules when abnormal results come in. All the discussion to date can be summed up by the admission that as yet there is no fairly simple, generally acceptable combining method that is practicable over a wide range of drilling conditions, ground conditions, and ore occurrence. The combining problem is important in evaluating drilling results at Chuquicamata. Recently a reappraisal has been made of recovery variables and their effect on assays, with the result that a new combining method is offered which fits average drilling conditions and is mathematically reasonable. It is simple in application, fundamentally correct, and an improvement over most combining methods. At Chuquicamata diamond drillholes are used to outline the grade and position of blocks of normal and marginal grade oxide, mixed, and sulphide ore. Most holes penetrate all classes of material (and waste), and it is important for mining programs as well as ore reserves to know almost precisely the soluble and insoluble copper content of mineralized ground. At present three classes of ore are mined and treated differently. For an orderly sequence of mining operations which can provide regular daily tonnages of all three ore types and keep grade at certain levels with minimum variation, ore type and its grade must be predicted. Diamond drilling plus geologic mapping and bench sampling are tools for prediction. And drilling data are largely used to calculate grade of material more than a few meters away from bench faces. The orebody at Chuquicamata" is criss-crossed by millions of barren or mineralized weak to fairly strong slips and fault fissures. Mineralization is diverse and encompasses many quartz and oxide or sulphide-bearing copper veins, as well as seams and disseminated grains. Copper occurs in oxide or sulphide minerals or mixtures of these two mineral types. Rock conditions vary from intensely seri-citized (soft and porous) through clay-altered ground to almost fresh granodiorite. The result is an orebody which offers many obstacles to good and consistent core recovery in diamond drilling. Recovery varies considerably in the several alteration zones, the various types of oxide and sulphide ores, and the position and inclination of the drillhole within the complex fracture pattern. As core recovery drops sludge samples must be used with core samples to calculate grade. Many years of drilling at Chuquicamata indicate that in good grade oxide zones and in the sulphide areas core recovery is good and the ground uniformly mineralized. Moderate loss of core, therefore, does not markedly affect grade calculations based on core assays. An early core-sludge combining method used core assays at face value as indicating grade down to 50 pct core recovery, but below this recovery percentage sludge samples were used and weighted according to the standard Longyear chart. This method apparently did not introduce serious errors, but it abruptly used sludge assays with high weighting factors representing 100 pct return irrespective of actual percentage of sludge recovered. Recent drilling activities have been directed toward outlining the marginal ore areas. The non-uniform mineralization and generally poorer core recovery in such ground indicated that a more exact core-sludge combining method was required to equate wide differences between core and sludge assays and recovery. In fringe ore areas at Chuquicamata core recovery averages about 50 pct, from a minimum of 10 to 15 pct to a maximum of 100 pct. Sludge recovery is likewise variable and averages perhaps 80 pct even though holes are cemented as water return falls off. In a homogeneously mineralized area cut by many slips and faults, with hard and soft ribs, loss of core is loss of ground which has a grade similar to that of core recovered, and core assays approximate true grade. In this case sludge samples need not be used. However, it would be unusual to know beforehand that an area is uniformly mineralized, and in fact this condition is probably uncommon. Generally the distribution of valuable minerals in the ground does not exactly compare with their recoverability in core. Thus, in the usual case, loss of core decreases the

Jan 1, 1956

By Alan K. Burton

Having estimated fleet size and productivity, the next and final step in this analysis is to introduce cost. Total cost of the system is a combination of capital, operating, and any other external costs that may be incurred. The last item, external costs, can be particularly important when comparing one size or type of truck fleet with another. Once costs are calculated, it is possible to compare systems on a total cost, cost per hour, cost per ton, or any other basis that is required. Methods used vary from simple owning and operating costs to sophisticated discounted cash flow analyses involving optional fleet replacements over the project life.

Jan 1, 1976

By Tell Ertl

THE term oil shale is defined. Foreign oil-shale developments are outlined briefly, The richest and most extensive oil-shale deposits in the United States lie in Western Colorado. The Bureau of Mines is opening two adjacent areas for underground mining in Western Colorado. From the first area the oil shale is mined selectively as required by the retorting plant. The second area is being developed as a unit of a full scale oil-shale mine to ascertain the most practical methods of mining oil shale and to determine costs. In both of these mining areas, research will be carried on to develop new and improved practices for mining oil shale.

Jan 1, 1947

By R. R. Klimpel Discussion by C. C. Harris, L. G. Austin

C. C. Harris (Columbia University) — The authors ' have presented relationships for the random division of a magnitude into r + l parts taking account of the order of selection. Interpreting their model in comminution terms, their Eqs. 7, 9 and 11 refer to the one-dimensional case; that is, to the subdivision of a long thin rod at random. It is assumed that the fragments may be indefinitely small. Real materials, however, are limited in the degree of subdivision which can be attained through the intervention of practical restrictions. One in fine grinding is the depletion of microflaws as material becomes finer. Another is the

Jan 1, 1969

By Thomas F. Donnelly

The Province of San Cristobal is situated on the south side of the island of Santo Domingo about 25 miles west of Santo Domingo city, the capital of the republic. The copper mineraliza,tion is found about 8 miles north of the town of San Cristobal,. in the section known as Bucaro Hill; in San Francisco Hill; on the Nigua River; and on the Jaina River. The district is reached by automobile from Santo Domingo city over an excellent road for 25 miles to San Cristobal, and then, by horseback 6 or 8 miles up the Nigua River to what is locally known as "Camp

Jan 1, 1916

By C. P. Parsons

CALIPER logging is a practice of measuring the variations in the diameter of the open hole in a well. This information is useful for many purposes, among which are: Determining the volumetric capacity of the portion of open hole to be cemented; the amount of gravel to be used for gravel-packing a well; the size and location of seat for a packer to be placed in a well; studying the effects of nitroglycerin shots; studying the effects of acidizing; studying the cause of drill-pipe failures; studying drilling technique; geological correlations. This is a new practice in oil fields and is eliminating considerable guesswork.

Jan 1, 1942

By S. L. Crouch

The application of a simple, two-dimensional computer technique for evaluating rock burst potential in cut-and-fill stopes was investigated. The principal purpose of the study was to compare the effectiveness of high-modulus fill and destressing as aids in the control of rock bursts. Two conclusions that are probably of general applicability are drawn. First, high-modulus fill is ineffective as a means of increasing the mine stiffness and acts to control rock bursts only if it supports enough load to reduce substantially the stresses on the unmined vein material. Second, destressing is most effective if it is performed when the vein material is near the point of failure.

Jan 1, 1975

By Kalanadh V. S. Sastry, D. W. Fuerstenau

A locked-cycle balling procedure has been designed to simulate continuous, closed- circuit balling drum operation in the laboratory. By this method, it is possible to simulate the surging in green-ball production rate and circulating load that is experienced in industrial balling circuits. Through an investigation of the effect of feed rate on the stability of the locked-cycle, balling process, it was found that higher feed rates lead to increasing instability in the process. An analysis of the experimental ,results is presented and the utility of the locked-cycle experimental technique for evaluating the stability of an industrial balling drum operation is discussed.

Jan 1, 1976

By John T. Galey

In all, 210 wellst were completed in southwestern Pennsylvania during 1939, which is 59 more than were drilled during the preceding year. Of this number, 137 were gas wells, which is a considerable increase from the 81 completed during 1938. Only 25 oil wells were completed, as against 28 in the previous year, and 28 dry holes, 6 more than last year. Shallow Development—Gas The increase in the number of gas wells is largely the result of a search for additional gas in Armstrong and Indiana Counties. Although a considerable number of wells were drilled in the latter county, the majority of them produced less than 100 M cu. ft. initial open flow. In Armstrong County, a new pool was opened in Cowanshannock township, producing from the third Bradford sand. Several large wells were obtained there, one of which produced an initial of nearly 12 million cu. ft., with a reservoir pressure of 1100 Ib. However, the productive area has been fairly well outlined by further drilling and shows that the pool covers only about 800 acres. The sand is 20 to 30 ft. thick, and probably has 5 to 6 ft. of pay. Thus this reserve is not an important one. Several wells were found in Menallen township, Fayette County, where the Big Injun sand has been having a play for the past three years, but the activity there is much less than a year ago. Oil On the west side of the Bradford sand gas pool, Cowanshannock township, Armstrong County, a well was drilled which flowed 100 bbl. the first day from the Tiona sand. The probability is that this well is a freak, as the Tiona has been drilled within a fairly close distance on every

Jan 1, 1940

By John T. Galey

In all, 210 wellst were completed in southwestern Pennsylvania during 1939, which is 59 more than were drilled during the preceding year. Of this number, 137 were gas wells, which is a considerable increase from the 81 completed during 1938. Only 25 oil wells were completed, as against 28 in the previous year, and 28 dry holes, 6 more than last year. Shallow Development—Gas The increase in the number of gas wells is largely the result of a search for additional gas in Armstrong and Indiana Counties. Although a considerable number of wells were drilled in the latter county, the majority of them produced less than 100 M cu. ft. initial open flow. In Armstrong County, a new pool was opened in Cowanshannock township, producing from the third Bradford sand. Several large wells were obtained there, one of which produced an initial of nearly 12 million cu. ft., with a reservoir pressure of 1100 Ib. However, the productive area has been fairly well outlined by further drilling and shows that the pool covers only about 800 acres. The sand is 20 to 30 ft. thick, and probably has 5 to 6 ft. of pay. Thus this reserve is not an important one. Several wells were found in Menallen township, Fayette County, where the Big Injun sand has been having a play for the past three years, but the activity there is much less than a year ago. Oil On the west side of the Bradford sand gas pool, Cowanshannock township, Armstrong County, a well was drilled which flowed 100 bbl. the first day from the Tiona sand. The probability is that this well is a freak, as the Tiona has been drilled within a fairly close distance on every

Jan 1, 1940

By R. W. Jigins

OPERATIONS of the Braden Copper Co. are in the Chilean Andes, southeast of Santiago. Most remote of the company communities is Sewell, a town of 12,000 people, 7000 ft above sea level at the junction of the Coya and Teniente rivers. Four miles down the Coya River is the smelter town of Caletones, elevation 4500 ft, with a population of about 3000. At the junction of the Coya and Cacha-poal rivers 24 miles below Sewell is Coya, elevation 2500 ft, location of the company's two hydroelectric plants. The principal accounting offices; steel, iron, and brass foundries; railroad shops; and major mechanical repair and fabricating shops are at Rancagua in the central Chilean valley, 44 miles from Sewell. The company owned narrow-gage railroad connects with the Chilean national railroad at Rancagua and above Coya is the only modern means of travel. El Teniente mine at Sewell is the largest underground copper mine in the world. The mill, at the portal of the main tunnel, has a capacity of 34,000 tons of crude ore daily, although production is lower during the winter beacause of electric power restrictions. When operating at capacity the mill produces 2000 tpd of copper concentrates and 32,000 tpd of tailings. Tonnage of tailings is equivalent to about 100,000 tons of pulp at normal water to solid ratio. Because the streams flowing from the Andes in the area of Braden's operations are used for irrigation, it is impossible to dump tailings directly to the river. The tailings pulp must be separated into solids and decanted clear water, and the clear water returned to the same water shed from which the original water was removed. Remaining sands and slime must be deposited in a reservoir where they will not cause trouble to farmers.

Jan 1, 1958

By G. G. Gunther, C. L. Boeke

THE zinc deleading processes at the Parral and Santa Barbara mills are described separately to provide a basis for comparison. Although the two procedures are fundamentally alike, there are some differences in application to meet specific conditions. The Asarco mill at Parral in the State of Chihuahua is a flotation plant treating 1450 metric tons per day of complex sulphide copper-lead-zinc ores. In general, the procedure used for selective separation of the minerals into a copper, a lead, and a zinc concentrate follows the usual practice. First, a bulk copper-lead concentrate is floated and then separated into a final copper and a final lead concentrate by the sulphurous acid process. A small tonnage of high grade gravity lead concentrate is tabled from the bulk copper-lead cleaner tailing before it is reground so as to avoid undesirable over-grinding and sliming of this portion of the lead mineral. Next, a zinc concentrate is floated from the lead rougher tailing and, as described later, is separated into a final lead product and a final zinc concentrate. Finally, the zinc rougher tailing receives a strong scavenger float to remove all recoverable middling values for regrinding and return to the flotation circuits. The slime is separated from the scavenger flotation product, and only the sand portion goes to the regrind section. The minerals in the separated slime are extremely resistant to all methods of selective separation and pose one of the major problems awaiting solution. Nevertheless, the slime shows a profit when mixed with either the lead or the zinc concentrate, except when metal prices are very low, so it is accepted as a final product for lack of a more profitable method of handling it. In contrast, the minerals in the reground sands of the

Jan 1, 1953

By John Henderson

FOR some time now the most commonly accepted description of liquid silicate structure has been the "discrete ion" theory, proposed originally by Bockris and owe.' This theory is that when certain metal oxides and silica are melted together, the continuous three dimensional silica lattice is broken down into large anionic groups, such as sheets, chains, and rings, to form a liquid containing these complex anions and simple cations. Each composition is characterized by "an equilibrium mixture of two or more of the discrete ions",' and increasing metal oxide content causes a decrease in ion size. The implication is, and this implication has received tacit approval from subsequent workers, that these anions are rigid structures and that once formed they are quite stable. The discrete ion theory has been found to fit the results of the great majority of structural studies, but in a few areas it is not entirely satisfactory. For example it does not explain clearly the effect of temperature on melt structure,3 nor does it allow for free oxygen ions over wide composition ranges, the occurrence of which has been postulated to explain sulfur4 and water5 solubility in liquid silicates. In lime-silica-alumina melts the discrete ion theory is even less satisfactory, and in particular the apparent difference in the mechanism of transport of calcium in electrical conduction8 and self-diffusion,' and the mechanism of the self-diffusion of oxygen8 are very difficult to explain on this basis. By looking at melt structure in a slightly different way, however, a model emerges that does not pose these problems. It has been suggested5" that at each composition in a liquid silicate, there is a distribution of anion sizes; thus the dominant anionic species might be Si3,O9 but as well as these anions the melt may contain say sis0:i anions. Decreasing silica content and increasing temperature are said9 to reduce the size of the dominant species. Taking this concept further, it is now suggested that these complexes are not the rigid, stable entities originally envisaged, but rather that they exist on a time-average basis. In this way large groups are continually decaying to smaller groups and small groups reforming to larger groups. The most complete transport data 8-10 available are for a melt containing 40 wt pct CaO, 40 wt pct SiO2, and 20 wt pct Al2O3. Recalculating this composition in terms of ion fractions and bearing in mind the relative sizes of the constituent ions, Table I, it seems reasonable to regard this liquid as almost close packed oxygens, containing the other ions interstitially, in which regions of local order exist. On this basis, all oxygen positions are equivalent and, since an oxygen is always adjacent to other oxygens, its diffusion occurs by successive small movements, in a cooperative manner, in accord with modern liquid theories." Silicon diffusion is much less favorable, firstly because there are fewer positions into which it can move and secondly, because it has the rather rigid restriction that it always tends to be co-ordinated with four oxygens. Silicon self-diffusion is therefore probably best regarded as being effected by the decay and reformation of anionic groups or, in other words, by the redistribution of regions of local order. Calcium self-diffusion should occur more readily than silicon, because its co-ordination requirements are not as stringent, but not as readily as oxygen, because there are fewer positions into which it can move. There is the further restriction that electrical neutrality must be maintained, hence calcium diffusion should be regarded as the process providing for electrical neutrality in the redistribution of regions of local order. That is, silicon and calcium self-diffusion occur, basically, by the same process. Aluminum self-diffusivity should be somewhere between calcium and silicon because, for reasons discussed elsewhere,' part of the aluminum is equivalent to calcium and part equivalent to silicon. Consider now self-diffusion as a rate process. The simplest equation is: D = Do exp (-E/RT) [I] This equation can be restated in much more explicit forms but neither the accuracy of the available data, nor the present state of knowledge of rate theory as applied to liquids justifies any degree of sophistication. Nevertheless the terms of Eq. [I] do have significance;12 Do is related, however loose this relationship may be, to the frequency with which reacting species are in favorable positions to diffuse, and E is an indication of the energy barrier that must be overcome to allow diffusion to proceed. For the 40 wt pct CaO, 40 wt pct SiO2, 20 wt pct Al2O3, melt, the apparent activation energies for self-diffusion of calcium, silicon, and aluminum are not significantly different from 70 kcal per mole of diffusate,' in agreement with the postulate that these elements diffuse by the same process. For oxygen self-diffusion E is about 85 kcal per mole,' again in agreement with the idea that oxygen is transported,

Jan 1, 1963

By D. Turnbull, J. H. Hollomon

IN the most general sense, solidification refers to the formation of crystalline material from either a gas or a liquid. However, in this symposium, only the formation of crystals from liquids will be considered. Thus, in our present study the problem of solidification becomes that of determining the mechanism of the transition from a liquid to a crystalline solid, and our particular interest is in the transformation of metals from the liquid to the crystalline state. Metal liquids are seldom composed of only one kind of atom. Often solute atoms are added intentionally to make an alloy, or they may be present accidently because of incomplete purification. Furthermore, there may be small solid particles of various other crystals present in the liquid. Such small crystals may persist if their melting points are higher than any temperature to which the liquid is heated. Prior to solidification, the liquid metal usually is poured from one container to another. The second container, called the mold, usually is at a lower temperature than the container, or the crucible, in which the metal is liquified. The first solid that forms upon contact with the mold generally is of a different composition than the liquid, usually containing fewer foreign atoms or dissolved impurities. Heat is evolved upon solidification and is dissipated through the surroundings. Tree-like crystals, or dendrites, grow from the container walls and branch into the melt, enriching the remaining liquid in impurity atoms. In some alloys the remaining, relatively impure liquid continues to solidify until no liquid remains and the branches of the original crystal are cemented together. In other alloys, particles with a second type of crystal structure appear before solidification is complete. First the nature of the liquid and the solid states will be considered and then the mechanism of the initial formation of crystals from pure and alloy liquids. Later the problem of the growth of these crystals and the complexities of solidification in a mold, where the temperatures are not uniform and the composition of the melt changes continuously, will be discussed. LIQUID AND CRYSTALLINE STATES A liquid is an aggregate of atoms arranged with no apparent regularity (over distances of more than a few atom diameters). Nevertheless, any given atom generally has a definite number of nearest neighbors at definite distances, and thus order may be said to exist over a short range. In a crystal, on the other hand, a regular array of atoms continues for many hundreds or thousands of atom diameters, and a long-range order is apparent. Thus, a crystal having long-range order can be reproduced by linear translations for 10-5 cm or more of some fundamental unit possessing short-range order. A crude illustration of the difference between liquid and crystalline structures can be obtained by comparing the possible arrangements of oranges in a crate. If the oranges are simply dumped into the crate haphazardly until it is filled, the oranges will represent a static model of a liquid.

Jan 1, 1951

By R. Henry Norton

The State of Oregon, although admitted into the Union February 14, 1859, is to-day, so far as any practical knowledge of its mineral resources is concerned, almost terra incognita, as compared with the other Western States and Territories. This is partly to be accounted for by the class of emigrants who first settled in the country ; they seem to have been, as they still are, a pastoral community, to whom the terms " mines " 'and " mining " are synonymous with bankruptcy and disaster, and who discourage all attempts to develop the mineral resources of the State. This statement is emphasized by the fact that Oregon has no State Geolo-

Jan 1, 1891

By L. M. Cathles, W. Joseph Schlitt

A mathematical model describing the leaching of low grade industrial waste dumps in developed and solved by finite difference techniques. The leaching behavior of model dumps of different sizes, permeabilities, and waste characteristics, under different operating conditions, is then investigated using the model. It is shown that leaching is fastest for dumps 50-100 feet (15-30m) high, that permeability is very important for the effective leaching of large dumps, that slow leach kinetics of the waste may not be detrimental, and that the leaching of large dumps may be facilitated if leach solutions are preferentially applied to the hot, near-edge portions of the dump.

Jan 1, 1980