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By Rocky Gunderson

When operations begin early next year at the $100-million Port Kembla coal loading facility in New South Wales, Australia, a sophisticated computer system, directed by two of Hewlett-Packard's HP1000 mini-computers, will perform a number of functions ranging from monitoring and reporting arrivals and departures to maintaining a library of events for up to one year. The three-processor HP1000 system configuration, valued at $215,000, is one of the largest single orders, in terms of dollar value, ever submitted in Australia for Hewlett-Packard's line of engineering and manufacturing minicomputers. The processors will be linked in a fully "backed-up'' configuration using Hewlett-Packard's fourth generation distributed networking system, DS/1000-IV, making the system one of the most sophisticated in Australia. Interfaced to a number of weightometers and programmable logic controllers (PLCs), the computers will be the heart of a coal inventory system for the facility which, in its first phase, will be capable of handling more than 13.5 Mt/a of coal. The computers will report on all plant functions, including conveyor belt and water spray operations. Specifically, the system will: • Monitor and report all coal movements related to road and rail deliveries, dispatches from ship loaders, and interstockpile transfers. •Maintain a ''running log" of all events pertaining to overall plant operations, such as conveyor starts and stops and anti-pollution spray activation. • Accept and maintain plant data entered manually or acquired automatically from various remote locations via either visual display units (VDUs) or PLCs. • Maintain complete historical records of all transactions and related plant events for up to one year. • Provide a facility for stockpile management, modeling-type programs, and training or other simulation activity. The facility was designed with the following characteristics in mind, based on an analysis of requirements for an automated coal loading operation that would optimize productivity and be flexible to change and grow with facility needs and changing technology: • Continuous Operations-Due to the need for continuous plant logging and control, the overall facility will be capable of uninterrupted, 24¬hour operation. • Handling Current and Projected Workloads-The facility will be capable of not only handling current workloads, but also any additional workloads required during future coal loading operations. • Expansion-The facility will be capable of expansion, for a minimum cost, at any time with minimum disruption in operations. • Single Source of Supply and Maintenance-The facility will be supplied and maintained by one supplier, through hardware and software maintenance support contracts. System Configuration Two HP1000 minicomputers, each with 500K bytes of main memory, will be front-ended by a pre-processor with 256K bytes of main memory. These three systems, active, backup, and front-end, will be interconnected into a triangular array, as shown in the accompanying diagram. Under normal conditions, the front-end will transmit plant status data to one of the main systems, whose primary function will be to monitor coal flow and maintain a relevant dialogue with operators. The front-end, therefore, handles all communication with the plant PLCs, effectively removing this overhead from the current plant monitor (either of the two main systems). The front-end pre-processes and post-processes all information flowing from the plant monitor to the PLCs and vice versa, thus effectively handling all communication between the two units. A second main system will handle reporting, administrative, and modeling functions, and will also be capable of performing both front-end and monitoring operations. Linking all

Jan 11, 1981

By W. L. Falke, A. A. Cochran

A basic problem in connection with manganese is to find economical ways to utilize domestic resources. As a part of its program to conserve domestic mineral resources and to reduce dependence of foreign sources, the Bureau of Mines had conducted extensive research to broaden manganese technology and to find economical, technically-feasible processes for recovering manganese from these low-grade resources. This paper analyzes some of the major difficulties which have been encountered in previously proposed processes for treating low-grade manganese ores and discusses a process which may circumvent many of these difficulties. Vast deposits of low-grade manganese ore exist in the U.S., yet domestic production is insignificant compared to the more than 2 million tons of ore imported yearly into this country.' A basic problem in connection with manganese is to find economical ways to utilize domestic resources.2 As a part of its program to conserve domestic mineral resources and to reduce dependence on foreign sources, the Bureau of Mines has conducted extensive research to broaden manganese technology and to find economical, technically feasible processes for recovering manganese from these low-grade resources. Although a great many processes have been proposed for treating domestic manganese resources, a failing of many of them is that too many steps are involved for economical operation.' Some of the major processes that have been proposed were reviewed by Norman and Kirby.3,4 Chlorination processes have become increasingly useful as a means of recovering valuable metals from raw materials, because the relative cost of chlorine (Cl2) and hydrogen chloride (HC1) has decreased through the years, and because the corrosion resistance of materials of construction has been improved. The objective of these processes is to form metal chlorides that are more volatile or water-soluble than are the corresponding oxides or sulfides; these compounds then can be separated from the gangue or other minerals by volatilization or leaching. Several chlorination processes for the recovery of manganese have been investigated by the Bureau of Mines5-8 and others; the major chlorination processes were reviewed in 1963, and their main features were compared.4 In all these processes, the simultaneous leaching or volatilization of iron chlorides with manganese chloride (MnCl2) is a problem, and the more iron that is present, generally the greater the problem. The leaching step is not highly selective, and the product of this step is, at best, suitable only for the production of ferromanganese or for further processing to remove iron. The product of the volatilization step is, in all cases, a mixture of iron and manganese chlorides that contains too much iron even for ferromanganese production, and must undergo additional processing steps for iron removal. THERMODYNAMIC CONSIDERATIONS The literature indicates that the desired chlorination reactions will proceed at practical rates at temperatures near 1200°K (927°C). Free energy changes at 1200°K for the critical steps of the reduction and chlorination reactions9-11 follow: Most of the research on chlorination processes for manganese recovery has been done with HC1 rather than with C12, for the following three main reasons: 1) Recycling of the chlorinating agent, which is economically vital, appeared to be easier for HC1

Jan 1, 1968

By Harold Margalin, Elmars Ence

M. H. Mueller, M. V. Nevitt, and H. W. Knott (Argonne National Laboratory)—We have reported1 the crystal structure of Ti2CU as body centered tetragonal with a. = 2.944A and co = 10.786A, six atoms per unit cell, and space groupl4/mmm(MoSi2 type). This cell is larger but definitely related to the unit cell reported by the authors. It permits definite positions for the titanium and copper atoms in the cell located as follows: 2 Cu (a) 0 0 0; 1/2 1/2 1/2 4 Ti (e) 0 0 z; 1/2 1/2 1/2 + z; 0 0z; 1/2 1/2 1/2- z where z = 0.341 The face centered tetragonal cell reported by the authors can be transformed to a small body centered tetragonal cell with flo (small bet) = ao(fct)v2 Both unit cells have the same CO. Neither the face centered tetragonal cell with four atoms per unit cell nor the small body centered tetragonal cell with its two atoms per unit cell lend themselves conveniently to an orderly arrangment of the 2 to 1 ratio of titanium to copper atoms. When the co of this small body centered cell is tripled, however, a large body centered tetragonal cell results with six atoms per unit cell as described above. We had also noted that the X-ray patterns of Ti2Cu indicated that the structure could be approximately described by the face centered tetragonal cell as reported by the authors; how- ever, there was evidence for a few additional very weak lines whose positions indicated a larger unit cell. Since the X-ray scattering amplitudes of titanium and copper are nearly equal, it is evident that it is difficult to distinguish their positions especially from X-ray powder patterns. It was therefore decided to look at a powder pattern obtained by neutron diffraction since the scattering amplitudes of the two elements are considerably different in magnitude and opposite in sign. The sample used for neutron diffraction was selected from a group of three alloys prepared by arc melting using crystal bar titanium of 99.92 pet nominal purity and 99.999 pet pure Cu. The alloys designated as A, B, and C had the following respective compositions: Ti-32.8 at. pet Cu, Ti-33.3 at. pet Cu, and Ti-33.8 at. pet Cu. On the basis of the melting losses incurred, it was established that the actual concentration of either component in any of the alloys probably did not deviate from the intended value by as much as 0.1 at. pet. The alloys were annealed for 5 days at 700°C in evacuated Vycor capsules and quenched in tap water. The metallographic structures of these alloys are shown in Fig. 1. None of the alloys contained sufficient second phase to permit its positive identification. Since alloy B contained the least amount of second phase and since its composition placed it closest to the ideal stoichiometry, it was chosen for the neutron diffraction study. The oxygen content of alloy B was found by chemical analysis to be 0.06 at. pet. This particular intermetallic compound is especially suitable for investigation with neutron diffraction since if the titanium and copper atoms have a random statistical distribution, a near null-matrix would result in which there would be practically no coherent scattering of neutrons, and hence no Bragg diffraction peaks. Such a condition could come

Jan 1, 1962

By W. H. Triplett

DISCUSSION John G. Barry (Mexico)—"The Nerinea limestone," writes W. H. Triplett, p. 584, "is ... medium-bedded (0.5 to 1.0 m)...." This is the Zuloaga formation. It has always been described as thick-bedded (1.0 to 3.0 m). "Whether the marls and shales were less amenable to replacement," he goes on to say, ". . . is a matter of opinion." In general, it is probable that the thinner-bedded formations were less competent than the nearby thicker-bedded, so that the latter produced pre-mineral breccias when all were subjected to flexing stresses. With reference to Mr. Triplett's analysis, p. 584, under Lower Cretaceous, in the Taraises and Cupido formations some few horizons should be considered as possibly containing hidden orebodies. In the Potrero camp, shown on the general map in the northwestern part of Fig. 4, p. 587, several mineralized beds are known; they are found throughout the whole thickness of the Cuesta del Cura formation. With regard to intrusive rocks, p. 585, the general map, Fig. 4, shows that the Potrero and Noche Buena camps are near a third, monzonitic intrusive. It seems doubtful that the Aranzazu was any hotter than the Providencia camp. However, the former does contain large deposits of exogenetic silicates, which in large part occur within the Zuloaga limestone, as shown by Bergeats in a stereogram. At the Aranzazti camp the La Caja formation was only slightly changed, whereas it was altered highly at the Providencia, possibly en-dogenetically. That the temperatures at the Providencia camp were relatively high at one time is also shown by the tourmaline veinlets and lenses in the monzonite and by the garnets on Temeroso peak; high temperature mineralization has possibly been eroded. Maps and sections show that the Concepcion del Oro —Temeroso stock occupies the original locus of some 40 kms of sedimentary rocks. The remains of large roof pendants at Carmen and Animas indicate that stoping was active. Whether all the sedimentaries were stoped and assimilated presents a lively problem, a correlative of which would be an estimate of the original constitution of the invading instrusive rock. As to extrusive rocks, see p. 585, there is also a large thick area of the extrusives left near Noche Buena in the "La Caja" intramontane valley of the La Caja range. This place gave its name to the range. The volcanics are also said to occur in considerable amount in the Mazapil valley.'' Turning now to the remarks under Regional Structure, p. 586, as partly indicated on the general map, Fig. 4, the Mazapil Mountains of the Mazapil region are a continuous, U-shaped range including a northern leg of some 24 km, a southern leg of similar length, and a uniting base which lies to the East and which measures some 10 km in a S 20 E direction. Thus the U-shaped mountains are continuous through a length of 58 km and also contain one continuous anticline, the adjacent valleys being synclinal. The northern leg of the U is called the La Caja range, the base is called the Concepcion del Oro range, and the southern leg the Santa Rosa range. The ranges in northeastern Zacatecas and southern Coahuila contain anticlines and the valleys contain synclines, which where thin are necessarily highly compressed isoclines with very steep dips. Many of the anticlines also have isoclinal components which are overturned, original mushroom tops being indicated. The anticline in the base of the U of the Mazapil Region, Concepcion del Oro range, is a recumbent anticline, over to the ENE, with a necessarily subjacent recumbent syncline.

Jan 1, 1953

By R. C. Edwards, T. G. Kirkland, Will Mitchell

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By R. C. Edwards, T. G. Kirkland, Will Mitchell

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By W. H. Triplett

DISCUSSION John G. Barry (Mexico)—"The Nerinea limestone," writes W. H. Triplett, p. 584, "is ... medium-bedded (0.5 to 1.0 m)...." This is the Zuloaga formation. It has always been described as thick-bedded (1.0 to 3.0 m). "Whether the marls and shales were less amenable to replacement," he goes on to say, ". . . is a matter of opinion." In general, it is probable that the thinner-bedded formations were less competent than the nearby thicker-bedded, so that the latter produced pre-mineral breccias when all were subjected to flexing stresses. With reference to Mr. Triplett's analysis, p. 584, under Lower Cretaceous, in the Taraises and Cupido formations some few horizons should be considered as possibly containing hidden orebodies. In the Potrero camp, shown on the general map in the northwestern part of Fig. 4, p. 587, several mineralized beds are known; they are found throughout the whole thickness of the Cuesta del Cura formation. With regard to intrusive rocks, p. 585, the general map, Fig. 4, shows that the Potrero and Noche Buena camps are near a third, monzonitic intrusive. It seems doubtful that the Aranzazu was any hotter than the Providencia camp. However, the former does contain large deposits of exogenetic silicates, which in large part occur within the Zuloaga limestone, as shown by Bergeats in a stereogram. At the Aranzazti camp the La Caja formation was only slightly changed, whereas it was altered highly at the Providencia, possibly en-dogenetically. That the temperatures at the Providencia camp were relatively high at one time is also shown by the tourmaline veinlets and lenses in the monzonite and by the garnets on Temeroso peak; high temperature mineralization has possibly been eroded. Maps and sections show that the Concepcion del Oro —Temeroso stock occupies the original locus of some 40 kms of sedimentary rocks. The remains of large roof pendants at Carmen and Animas indicate that stoping was active. Whether all the sedimentaries were stoped and assimilated presents a lively problem, a correlative of which would be an estimate of the original constitution of the invading instrusive rock. As to extrusive rocks, see p. 585, there is also a large thick area of the extrusives left near Noche Buena in the "La Caja" intramontane valley of the La Caja range. This place gave its name to the range. The volcanics are also said to occur in considerable amount in the Mazapil valley.'' Turning now to the remarks under Regional Structure, p. 586, as partly indicated on the general map, Fig. 4, the Mazapil Mountains of the Mazapil region are a continuous, U-shaped range including a northern leg of some 24 km, a southern leg of similar length, and a uniting base which lies to the East and which measures some 10 km in a S 20 E direction. Thus the U-shaped mountains are continuous through a length of 58 km and also contain one continuous anticline, the adjacent valleys being synclinal. The northern leg of the U is called the La Caja range, the base is called the Concepcion del Oro range, and the southern leg the Santa Rosa range. The ranges in northeastern Zacatecas and southern Coahuila contain anticlines and the valleys contain synclines, which where thin are necessarily highly compressed isoclines with very steep dips. Many of the anticlines also have isoclinal components which are overturned, original mushroom tops being indicated. The anticline in the base of the U of the Mazapil Region, Concepcion del Oro range, is a recumbent anticline, over to the ENE, with a necessarily subjacent recumbent syncline.

Jan 1, 1953

By A. Rosenzweig, M. M. Karnowsky

The generalized binary diagram of the Au-Sn-Pb system is presented. The chief characteristic is the pseudobinary between lead and AuSn. On the high-gold side of the pseudobinary, four reaction isotherms were found at 363°, 274°, 225°, and 204°C. On the low-gold side of the pseudobinary, three reaction isotherms at 277°, 205°, and 176°C were found. A vertical section from the eutectic solder composition to gold was developed from the isothermal sections generated. METALLURGICAL problems arising from the intentional or accidental formation of Au-Sn-Pb alloys in soldering applications have indicated a need for information on the phase equilibria in this ternary system. The literature indicates that a complete diagram presenting this information has not been prepared. Even the recent excellent work by prince' entitled "The Au-Pb-Sn Ternary System" actually extends only to the AuSn or 50 at. pct Au range. There are more events in the 50 to 100 at. pct Au than in the less than 50 at. pct range. This article by Prince furnished much guidance in preparing this report. Initial efforts with the isopleth from the eutectic solder composition 61.9 wt pct Sn-38.1 wt pct Pb and 0 to 90 wt pct Au indicated that extensions of the work to the whole of the diagram (even if in less detail) were necessary to understand this system. In undertaking the construction of such a diagram, this study used differential thermal analyses (DTA) and metallography with verification and identification of phases by electron microprobe analyses. EXPERIMENTAL PROCEDURE The procedure began with the development of a cell which would be unreactive with lead, tin, and gold, retain its strength to 1100°C, exclude oxygen, and also furnish a suitable well for insertion and protection of a thermocouple to monitor temperature changes in the material. Such a cell would make it possible to reuse the sample should a repetition of determination be desired. Fig. 1 is a photograph of two such quartz cells as finally developed. The specimens have been run many times. The cell is made up of an inner sample bucket and an outer capsule with a thermocouple well which the sample surrounds. The DTA unit was an assembly devised from available instrumentation. The sample holder was a 3-in.-long cylinder of nickel, 2 1/2 in. in diam, fitted securely in a vertical tube furnace. Two holes 7/8 in. in diam were symmetrically located to receive the quartz cells. A dummy cell with nickel powder was used as the reference. The absolute temperature was recorded with a Pt/Pt 10-pct Rh thermocouple with an automatic Kaye ice bath compensator. (Calibration indicated the compensator to be accurate to better than ±0.5° over the 100° to 1000°C range.) The absolute temperature was measured in the sample and recorded on an L&N mul-tirange recorder with a full-scale display of 1 mv and automatic ranging of 99 mv. The differential signal between the sample temperature and the reference temperature was amplified with an L&N microvolt amplifier and recorded on an L&N Speedomax G recorder. Since two charts were obtained, the chart speeds were equalized and each event was noted on the differential chart, and the temperature at that time was obtained from the temperature record.

Jan 1, 1969

By R. C. Edwards, Will Mitchell, T. G. Kirkland

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By R. C. Edwards, Will Mitchell, T. G. Kirkland

A scrubber of new design has been invented for the beneficiation of glass and chemical sands. The machine is described and its capacity and metallurgical efficiency compared with the performance of other sand cleaning devices. THREE years ago, when the Process Research Laboratory at Allis-Chalmers Manufacturing Co. sought a remedy for the increasing cost of disposing of great quantities of spent sands from foundries, R. C. Edwards developed a continuous inclined paddle scrubber which removed a substantial amount of the charred coatings from the grains with a surprisingly small amount of degradation of the sand itself.' The satisfactory results obtained on foundry sand, using this piece of equipment in conjunction with classification, suggested that a similar procedure could probably be applied to the cleaning of other sands, such as those required by the glass and chemical industries. This paper describes the flowsheet and the results of tests performed on impure natural sands. The four materials tested were silica sands obtained from widely separated geographical areas. Three of the sands tested contained various amounts of iron oxide and the fourth an excessive amount of calcareous bonding material. The iron oxide was present predominantly in the form of grain coatings, which were scuffed by the paddle scrubber, and in all cases a product was manufactured which would meet one of the sets of specifications imposed by the glass industry.' The Edwards paddle scrubber used in these tests consists of a rubber-lined steel cylinder 18 in. in diam and 48 in. long, mounted on a frame so that its slope can be varied 5" to 35" from horizontal. This variable slope permits controlled flow of material through the machine. Rubber paddles rotate inside the cylinder and are secured to a center shaft 2 in. square, which extends through the entire length and is turned at both ends so that it can be supported by sleeve bearings centrally fixed in the end plates of the shell. At the upper or feed end it extends through the bearing on the end plate into a sheave, to which the motor drive is connected by means of Texrope transmission. The four sets of paddles mounted on the shaft, two in each set, are made of rubber belting, 1/4 in. thick, 9 in. wide, and 16 in. long. The two paddles in each set are secured on opposite sides of the square shaft, but in the same longitudinal position along the shaft axis. In other words, the two paddles in the set operate 180" apart. Each successive set is mounted 90" from the preceding one. One end of each paddle is bolted to the square shaft and the other weighted by a steel bar, lx1 1hx9 in. long, which is attached about Yz in. from the outer tip of the paddle. By centrifugal force at proper shaft speed, these weights force the paddle surface to rub smoothly against the rubber lining of the cylinder. During operation, the sand forces its way between the paddle face and the surface of the cylinder, providing the necessary abrading action as the pulp flows by gravity down the length of the cylinder. The capacity of the machine is limited by the amount and the hardness of the coating material and by the quality of product desired. The degree of scrubbing is governed by the retention time in the scrubber, which in turn may be changed by varying the feed rate, water content of the feed, and

Jan 1, 1953

By J. S. Aronofsky, J. P. Heller

This paper presents a mathematical analysis of the fluid mixing which occurs during flow through porous media. The analysis is based on the well-known diffusion equation with mass transfer term. It is pointed out that the use of this equation is justified by a single general assumption which does not specify any particular mechanism for the mixing process. Formulas are given for tracer fluid concentrations for two different boundary and initial conditions. Calculated numerical values compare closely with some published results of experiments in which no viscosity or density difference existed between displaced and displacing fluids. INTRODUCTION A phenomenological theory of the mixing and diffusion process is presented for the flow in a porous medium of two miscible liquids of equal viscosity and density. This is based on the classical diffusion equation which has been used to explain such processes as Brownian motion and heat conduction. More recently, this same differential equation has been derived (in a manner analogous to the work in this paper) for the problem of heat transfer during fluid flow in porous media.' The purpose of this paper is to show that the available experimental evidence justifies the use of this simple diffusion equation. The recent data of Koch and Slobod' and of von Rosenberg' are used to test this diffusion model by comparing concentration curve shapes and calculating "effective diffusion coefficients". The same differential equation for miscible displacement, as well as some experimental confirmation, has been published in Japan by K. Yuhara,4 and recently in this country by Day.5 Yuhara arrives at his con-clusions through analogy with turbulent diffusion, de- scribing the microscopic velocities in terms of "eddy motion" — although unlike the ordinary eddies of hydraulics. Day proceeds directly from Scheidegger's6 concept of the dispersion of velocities. After a very clear description of Scheidegger's theory, he applies it to calculate the motion first of a drop and then of a thin layer of brine into a bed of sand. For the latter case, Day derives and solves the same differential equation as Eq. 1. In our presentation no detailed assumptions are made as to the mechanism of displacement. In fact, several microscopic processes may well be involved—the only general assumption made is that their cumulative microscopic effect may be described by the addition of a second-order term to the differentirl equation of continuity. DERIVATION OF EQUATIONS Let us consider the displacement in one spatial dimension of Liquid A from a sample of porous media by Liquid B. A and B are assumed completely miscible and of the same viscosity and density. Both fluids are also assumed to be incompressible and it is specified that the total flow is the same for any cross section. It is evident that the rate at which Liquid B would be carried across such a section would be given by the product of the fluid velocity, u, and the concentration, C, of B in the fluid mixture assuming the constancy of u and C over the cross section. Since the volume of Fluid B is conserved, this leads to the one-way wave equation - u ?C/?x = ? ?C/?l where x increases in the direction of u—the "Darcy velocity", Q/A (or velocity flux per unit area), and 9 is the fractional porosity. Consider the case where fluid B is displacing Fluid A and the concentration distribution is represented by a step funcLon (with a vertical discontinuity). The solution of Eq. 1 predicts no mixing

Jan 1, 1958

By David Cornell, Yusuf K. Sukkar

The fundamental differential equation for fluid flow has been rearranged, integrated numerically for natural gases, and the results presented in tabular and graphical form suitable for the direct calculation of vertical gas flow problems. The integration is rigorous except for the use of a constant average temperature. Both static and flowing bottom-hole pressures may be calculated by (his method without resort to trial and error procedures. INTRODUCTION Bottom-hole pressures can be determined either by direct measurement with a bottom-hole pressure gauge or by calculation. Since measurement with a bottom-hole pressure gauge is costly, calculation of bottom-hole pressures is to be preferred when possible. Calculation involves knowledge of the wellhead pressure, the properties of the gas, the depth of the well, the flow rate, the temperature of the gas, and the size of the flow line. These quantities must be combined in a suitable equation for vertical flow. Several equations have been developed for the purpose of calculating gas flow through both horizontal and vertical pipes1,2,4,5,6,7,8,9 These equations all have the basic differential equation for flow as their starting point. They differ only in the assumptions which are made in order to simplify the integration step. The bottom-hole pressures calculated using any of the better known equations are in good agreement with bottom-hole pressures calculated by a more rigorous but considerably more tedious stepwise integration method. All of these methods, however. require a trial and error solution for the calculation of flowing bottom-hole pressures. The purpose of this paper is to present a method which provides a direct solution of static and flowing bottom-hole pressures without use of a trial and error solution. At the same time the method involves fewer simplifying assumptions than any of the previous methods used, except lor the lengthy stepwise method. The derivation of any practical flow equation begins with the basic differential equation which is rigorous for almost all flow problems. The five terms in Equation 1 represent the potential energy, the potential energy of position, the kinetic energy, the shaft work done, and the energy losses due to motion of the fluid respectively. Equation 1 reduces to Bernoulli's theorem for the case of horizontal flow of an incompressible fluid, no work done, and no energy losses due to friction. In the present case, the term (v) (dv)/(g,.) can be shown to be quite negligible compared to the other terms, and is therefore neglected in the derivation. The term — dws includes work done by a pump or turbine in the system and is zero in this case. The frictional losses in the pipe are included in the term — d(lw) which is expressed in terms of the Moody friction factor as in Equation 2. (f) (v2) (dL) d(lW) = (f)(V2) (dl)/(2) (g) (d) The specific volume of a gas can be expressed in terms of the temperature, pressure, molecular weight, and compressibility factor. The velocity of the gas can be expressed in terms of the weight rate of flow, the specific volume, and the dimensions of the pipe.The reduced pressure Pr is to be used because z can be expressed as a function of Pr and Tr alone for most natural gases:

Jan 1, 1956

By Dale S. Preece

INTRODUCTION The Waste Isolation Pilot Plant (WIPP) is a research facility located in Southeastern New Mexico. The WIPP is being developed by the U. S. Department of Energy (DOE) to demonstrate the safe storage of defense-related radioactive wastes in bedded salt. Since the WIPP is a research facility. a number of large-scale in situ experiments have been planned and are currently under construction (Munson, 1983). These experiments were designed to study various aspects of nuclear waste storage in bedded salt such as mechanical responses and creep closures of drifts, thermal response due to heated canisters, and thermally-induced fluid migration. One purpose of the WIPP is to develop and demonstrate a general predictive structural analysis capability for a bedded salt repository. One set of experiments, called the Geomechanical Evaluation (Room G), is heavily instrumented to study the creep around several rooms with different configuration. A Layout of the Geomechanical Evaluation room is shown in Figure 1. One portion of it consists of wedge shaped pillar where two drifts intersect at an angle of 7.5 degrees. The pillar can be seen at the bottom of Figure 1. One purpose of the wedge pillar experiment is to study progressive pillar failure in the tip region. Another is to determine how nonuniform pillar thickness affects creep closure of the drifts. An important aspect of the experiments is the correlation between experimental data and corresponding pretest finite element analyses of the site. The finite element simulations serve two purposes. First, the computer simulations aid in understanding the experimental results by providing calculated stress. strain and displacement fields that cannot be measured directly. Second, comparison of calculated and measured drift closures serves as a method for validating or improving the finite element models and the constitutive models employed. The wedge pillar geometry required a 3-D finite element creep calculation. Results from these 3-D calculations will be presented in this paper. A 2-D plane strain double drift model which approximates the wedge pillar geometry at a slice perpendicular to the drift has also been performed and will be compared to the 3-D results in this paper. FINITE ELEMENT COMPUTER PROGRAMS Two finite element computer programs. JAC (Biffle, 1984) and JAC3D (Biffle. 1986), were used. JAC is a 2-D finite element program developed for quasi-static analysis of non-linear solids. It employs the conjugate gradient iterative technique to obtain a solution Spatial integration is performed using a single gauss point in each four node quadrilateral element. An hourglass viscosity technique is used to control the zero energy modes that occur with single point integration. The single point integration combined with the explicit nature of the program and exploitation of CRAY-1 computer architecture results in very efficient execution J4: results were compared to results obtained with eight other 2-D structural creep computer codes in the second WIPP benchmark exercise (Morgan, 1981) JAC3D was derived from JAC to treat 3-D finite element models and has many of the same characteristics including single point integration and hourglass viscosity. These characteristics have made 3-D creep analyses more reasonable by significantly reducing computation time The first exercise of the creep capability

Jan 1, 1986

By S. A. Falconer

This paper deals with the practical application of the Dutch State Mines cyclone separator for fine-coal cleaning. The more important operating variables are discussed, and results of a number of continuous-scale tests on various fine coals with a 6-in. cyclone separator are given. Descriptions of the Cyanamid Pilot Plant and the 50 tph Cyclone Separator Plant at the Emma Washery in Holland are included. TODAY, as never before, the coal producers of this country are faced with the necessity for obtaining maximum yield of marketable coal of all sizes down to and including the finest. Particular interest is being displayed in a more effective means of recovering the finer sizes which in the past so often have been discarded to waste or treated by inadequate equipment. The reasons for this interest are easily understood. Rising mining costs have resulted in the increased use of mechanized equipment underground for full seam mining. This, in turn, means more refuse in the run-of-mine and also more fines in the feed going to the cleaning plant. Stringent antistream pollution laws now make it necessary at least to impound the fines, and usually it is a paying proposition to recover the coal from these fines before sending them to the storage area. Finally, the market and price of the finer sizes has increased, but at the same time specifications have tightened and the demand is for higher grade products. During the past four years a great deal of publicity has been given to this new process and apparatus for treating fine sizes of ores as well as coal. Excellent discussions of the apparatus itself and the theoretical aspects of the basic operating principles involved in its operation have been presented by M. G. Driessen,1-3 M. R. Geer and H. F. Yancey,4,5 J. W. Hyer6 and more recently by D. A. Dahlstrom8 as well as others.7,9-11 In view of this, it has been decided to include in the present paper only a brief description of the cyclone separator and its operating principles. This paper deals primarily with some of the more practical commercial aspects of the process and equipment required for its operation; the variables in the operation of the cyclone, their effect and means for control; and the results obtained with a 6-in. semicommercial size separator on a variety of fine coals, bituminous and anthracite, using finely ground magnetite for the separating medium solids. A description is included of the commercial-size cyclone separator pilot plant installed by the Dutch State Mines at its Emma mine in Holland, where finely ground slate is utilized as medium solids. Field of Application: The Dutch State Mines cyclone separator operates most efficiently on ores and coal in the size range of % in. to about 48 mesh. On some types of feed, effective separation can be made on sizes as small as 100 mesh, but up to the present, this may be considered exceptional. Depending upon the type of medium used, the specific gravities of separation can be varied within wide limits, up to as high as 2.70, if magnetite is used as medium. Early Development: The development of the Dutch State Mines cyclone as a separator, or cleaner, was the outgrowth of early experimental work by Dries-sen' and others in Holland, just prior to World War 11, in the course of which was discovered the effectiveness of the cyclone in reclaiming and thickening the dense medium used in the "Loess" process of

Jan 1, 1951

By S. A. Falconer

This paper deals with the practical application of the Dutch State Mines cyclone separator for fine-coal cleaning. The more important operating variables are discussed, and results of a number of continuous-scale tests on various fine coals with a 6-in. cyclone separator are given. Descriptions of the Cyanamid Pilot Plant and the 50 tph Cyclone Separator Plant at the Emma Washery in Holland are included. TODAY, as never before, the coal producers of this country are faced with the necessity for obtaining maximum yield of marketable coal of all sizes down to and including the finest. Particular interest is being displayed in a more effective means of recovering the finer sizes which in the past so often have been discarded to waste or treated by inadequate equipment. The reasons for this interest are easily understood. Rising mining costs have resulted in the increased use of mechanized equipment underground for full seam mining. This, in turn, means more refuse in the run-of-mine and also more fines in the feed going to the cleaning plant. Stringent antistream pollution laws now make it necessary at least to impound the fines, and usually it is a paying proposition to recover the coal from these fines before sending them to the storage area. Finally, the market and price of the finer sizes has increased, but at the same time specifications have tightened and the demand is for higher grade products. During the past four years a great deal of publicity has been given to this new process and apparatus for treating fine sizes of ores as well as coal. Excellent discussions of the apparatus itself and the theoretical aspects of the basic operating principles involved in its operation have been presented by M. G. Driessen,1-3 M. R. Geer and H. F. Yancey,4,5 J. W. Hyer6 and more recently by D. A. Dahlstrom8 as well as others.7,9-11 In view of this, it has been decided to include in the present paper only a brief description of the cyclone separator and its operating principles. This paper deals primarily with some of the more practical commercial aspects of the process and equipment required for its operation; the variables in the operation of the cyclone, their effect and means for control; and the results obtained with a 6-in. semicommercial size separator on a variety of fine coals, bituminous and anthracite, using finely ground magnetite for the separating medium solids. A description is included of the commercial-size cyclone separator pilot plant installed by the Dutch State Mines at its Emma mine in Holland, where finely ground slate is utilized as medium solids. Field of Application: The Dutch State Mines cyclone separator operates most efficiently on ores and coal in the size range of % in. to about 48 mesh. On some types of feed, effective separation can be made on sizes as small as 100 mesh, but up to the present, this may be considered exceptional. Depending upon the type of medium used, the specific gravities of separation can be varied within wide limits, up to as high as 2.70, if magnetite is used as medium. Early Development: The development of the Dutch State Mines cyclone as a separator, or cleaner, was the outgrowth of early experimental work by Dries-sen' and others in Holland, just prior to World War 11, in the course of which was discovered the effectiveness of the cyclone in reclaiming and thickening the dense medium used in the "Loess" process of

Jan 1, 1951

By F. W. Smith

Within the Northern Pennine Orefield, England, intersections of fluorite veins are generally richest in ore, and their recognition enables investigation of the ore potential of the intersecting bodies. A method, based on routine analysis of fluorite for yttrium, has been used with success in mines of the district to predict the location of intersections in partially developed fluorite-bearing ore bodies. This short note is intended as a brief introduction; the details of United Kingdom applications are to be published at a later date. Geology The Northern Pennine Orefield is the second largest fluorspar-producing region in Britain. The associated minerals galena, sphalerite, chalcopyrite, barite, and witherite have all been extensively worked in the past but, with the exception of a single barytes mine, all of these operations are now abandoned. Galena alone is recovered as a byproduct of fluorspar beneficiation. At the present time two companies are producing about 38,000 tpy of fluorspar from high grade vein ore bodies, and a number of concerns are actively prospecting in old lead workings. The mineralization is localized in veins and replacement flats within favorable lithologies in an alternating limestone, sandstone, shale sequence. Since the thickness of individual beds is not great (generally less than 20 m), vein bodies tend to be ribbon-shaped, with length greatly exceeding width and height (Dunham). They thus present difficult exploration targets for conventional diamond drilling techniques. Ore grade varies greatly along the veins, which can change rapidly in character from mineral stained fault gouge to massive mineral bodies up to 10 m in width. It has been known for a very long time that the richer ore bodies tend to be associated with vein intersections in favorable beds. It is assumed that vein fissure intersections provided vertical channelways through which hot ore-bearing brines were able to ascend from deeper cavities in the sequence. Thus investigation of intersections at other horizons generally yields further ore bodies. Technique The relative concentration of yttrium in fluorite is a good indicator, in the North Pennines, of the direction of flow of mineralizing fluids. In all of the ore bodies so far studied in detail in this region, the concentration of yttrium increases markedly in the direction of vein intersections that are believed to have acted as feeder zones to the remainder of the bodies. The reason for this relationship is probably connected with the degradation of soluble rare earths-fluoride complexes by the changes in various physicochemical parameters of the hydrothermal fluids as they entered the ore body cavities from the feeder zones. The simultaneous release of free fluoride ions also contributed to the increased ore values in these intersection zones. Yttrium values in fluorite taken from a longitudinal traverse of the western ore body, Redburn mine, Rookhope, are plotted in [Fig. 1.] The position of a major vein branch and three cross veins are clearly indicated by yttrium maxima. The cross veins themselves contain up to 2.5 m of low grade fluorspar ore. The exact positions of B and C were unknown previous to the geochemical survey. Similar results have been obtained for several other veins and ore bodies in the area. Clean fluorspar was collected as bulk samples from within 300 mm of the vein wall and, since the vein infilling is simple, with mineral growth from the walls inward, analyses of the samples are considered comparable along the length of the ore body. A number of samples taken from the center of the vein show a

Jan 1, 1975

By R. W. Olson, A. L. Benham

Viscous fingering was studied as it occurred in an open Hele-Shaw model (1 ft x 4 ft x 1/16 in.); it was also studied in the same model packed with 80-mesh glass beads during miscible displacements under unfavorable viscosity ratio conditions. It was determined that the lengths of 20 elements of the front were distributed normally around the average position of the combined elements under all conditions. Represented by the normal distribution, the length of the viscous fingers grew linearly with the distance traveled from the point of finger formation, increased with displacement velocity, and increased with increasing mobility ratio. Results obtained during the first few inches of displacement were of little or no use in predicting finger growth and/or finger length throughout the 4-ft model since the point of finger initiation cannot be predicted and would often occur several inches from the point of injection with either positive or apparent negative coordinates. The fingering occurring during a miscible slug displacement was much greater than would be predicted based upon the actual mobility ratios between in-place fluid and slug and between slug and following fluid, using the results described above for miscible displacement in the absence of a slug. Many of the experiments in the packed model showed that the rate of growth of the viscous fingers was diminishing toward the end of the displacement in the 4-ft long model, indicating that microscopic mixing, such as diffusion or dispersion, was decreasing the viscous fingering effect. INTRODUCTION Viscous fingering is a manifestation of a finger-shaped interface between displaced and displacing fluids occurring during typical miscible displacement projects for oil recovery. Its cause may be traced to the instability of a viscous fluid being displaced by a more mobile fluid. Viscous fingering takes on important significance in the miscible slug process 1.2 where it may be a dominant factor in determining minimum slug size. Experimental1-3 and theoretical4-7 studies of viscous fingering have been made by other investi- gators. However, an exhaustive study of the variables affecting viscous fingering had not been made. The present study was undertaken in an effort to determine the effects of some of the more obvious variables—such as mobility ratio, displacement velocity, distance displaced, and packing—upon viscous finger length and growth in a small laboratory model. It does not necessarily follow that conclusions reached by studying the results of this model study may be applied to the field. Future studies would have to evaluate the effect of model size on extension of these results. EQUIPMENT AND EXPERIMENTAL PROCEDURES A flow diagram for the equipment used in this study is shown in Fig. 1. Basically, it consisted of (1) a constant-rate pump for the injection of displacing fluid into (2) a model made up of two flat transparent plates spaced a small distance apart and (3) provisions for the production of fluid into a calibrated graduate. The displacing fluid contained a dye which allowed visual observation of finger spacing and length. The pump performed at rates from 2.5 cc/hr to 480 cc/hr, while pumping a light mineral oil into a bladder in a closed glass container holding the displacing fluid. The model was made up of two plates of 1-in. Plexiglass with dimensions 1-ft wide by 4-ft long. These were spaced 1/16 in. apart so that the internal model dimensions were 1-ft wide x 4-ft long x 1/16

By Howard Martens

HARDNESS behavior of commercial titanium alloys following various heat treating processes has been studied for some time. However, the hardness of such alloys following a definite measured cooling rate from the single phase region has not been reported. Therefore, a series of experiments was conducted on the three alloys: 150A, 130A, and 130B. The manufacturers gave the following compositions for these alloys: Ti 150A: 0.045 pct C, 0.069 pctN, 1.32 pct Fe, and 2.68 pct Cr; RC 130A: 0.13 pct C, and 7.9 pct Mn: RC 130B: 0.11 pct C, 3.5 pct Mn, and 3.2 pct Al. A method similar to that described by Greninger was used, and the cooling rate varied from 6" to 9200°F per sec. The cooling curves were recorded on a Speedomax or a recording oscillograph, and the hardness was measured on a Tukon hardness tester using a Knoop indentor with a 500 g load. The curves of Fig. 1 show the variation of the Knoop hardness with the cooling rate for these three alloys. The alloys 150A and 130B behave in a similar manner, as shown by the curves. In neither of these alloys was it; possible to retain the high temperature ß phases, even by the most rapid cooling rate. The slow cooling rates produced a typical ß structure in these two alloys. As the cooling was increased, these phalses became more finely dispersed and the hardness increased. This increase in hardness with cooling rate continued for the 150A alloy until a rate of approximately 1000°F per sec was reached. At this cooling rate, the structure showed the result of the dif-fusionless transformation of the ß solid solution to the supersaturated a solid solution which is referred as a'. At higher cooling rates, the structure showed no change and the hardness showed no great change, as indicated by the dotted portion of the curve in Fig. 1. The increase in hardness with cooling rate for the 130B alloy continued until a rate of approximately 300°F per sec was reached. At this rate the structure was produced and the maximum hardness was reached. Higher cooling rates caused no marked changes in hardness or structure. The alloy 130A behaved in a different manner, as shown by the curve of Fig. 1. The structure produced by the slow cooling rates was a typical a+ß structure, which became finely dispersed as the cooling rate increased. The hardness increased with cooling rate and reached a maximum at approximately 120°F per sec. At this maximum hardness, the structure showed the result of the decomposition of the ß phase into the ß + phase structure. As the cooling rate was increased, smaller amounts of the at phase were formed and the alloy became softer. For cooling rates of 3000°F per sec and above, the high temperature ß phase was retained and the hardness of the alloy did not change markedly. Acknowledgment This paper presents the results of one phase of research carried out at the Jet Propulsion Laboratory, California Institute of Technology, under Contract No. DA-04-495-0rd 18, sponsored by the Dept. of the Army, Ordnance Corps.

Jan 1, 1957

By Eivind B. Jensen

Introduction Though temporarily shutdown due to the depressed economy, the Henderson mine is a large underground molybdenite producer, using a continuous panel-caving mining system. The location of the ore body requires access from the surface through three shafts, the collars of which are located about 3.15 km (10,350 ft) above sea level. At this altitude, normal temperatures are such that heat must be added to intake-ventilation air for about half of each year. This prevents ice from building up in the intake shafts, waterlines from freezing underground, employee discomfort, and equipment-related problems such as pneumatic rock drill icing and cold starting of diesel engines. Because of topographic and environmental constraints, the Henderson mill and concentrator are located 24 km (15 miles) from the mine, on the west side of the Continental Divide. All ore from the mine is transported to the mill by an electrified rail haulage system, with 15.4 km (9.6 miles) of the trip being underground through a double-track tunnel. A shaft at the midpoint of the tunnel exhausts air drawn from the mine and from the portal that is at an elevation of 2.7 km (8,950 ft). Intake air at the portal must be heated for part of the year to prevent ice buildup. Mine Ventilation System The mine ventilation system uses three sources of fresh air and two exhausts. The primary freshair intake is No. 3 shaft. It is concrete lined to a diameter of 7 m (23 ft) for a total depth of 698 m (2,291 ft). An airflow of 755 m3/s (1.6 mil¬lion cfm) is supplied to this shaft by two vane-axial 932-kW (1,250-hp) fans on the surface. The air is distributed within the mine by means of fresh-air distribution levels connected to production and development levels by raises. A secondary source of fresh air to the mine is provided by No. 2 shaft, the primary man and material access. This shaft is concrete lined to a diameter of 8.5 m (28 ft) for a total depth of 946 m (3,105 ft). Downcast airflow is induced by a pressure differential produced by the fans in the system such that net intake and exhaust volumes are equal. The No. 2 shaft flow of 165 m3/s (350,000 cfm) enters the mine network from the production and haulage levels. Exhaust air is moved from the various working levels through connecting raises to exhaust col¬lection levels that return it to No. 1 shaft. No. 1 shaft is the primary mine exhaust. It is concrete lined to a diameter of 7 m (23 ft) for a total depth of 788 m (2,584 ft). Three vane-axial 932-kW (1,250-hp) exhaust fans are located at the collar of No. 1 shaft, producing a flow of 850 m3/s (1.8 million cfm). In addition, an airflow of 71 m3/s (150,000 cfm) from the intake distribution level is exhausted via the haulage tunnel, by means of No. 4 shaft located midway in the tunnel. The main features of the mine ventilation network are summarized in Fig. 1. Ventilation Air-Heat Plants No. 3 Shaft In the early design phase of No. 3 shaft, it was recognized that air heating would be required to maintain underground temperatures above 0°C (32°F) during cold winters in the Colorado mountains. When the air reaches the mining

Jan 1, 1983

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

It is imperative that the hydrogeologic frame- work and the ground water flow systems of a Potential mine area be investigated prior to initiation of mining. Such investigations are necessary so that potential mine water inflow problems can be anticipated and controlled. Obvious difficulties can arise if large inflows of ground water occur in underground workings. Costs of water removal and disposal, if not foreseen and included in economic planning, can severely limit or eliminate profit from a mining operation. Safety problems associated with large inflows are an additional potential problem. The inflow of water into a mine necessarily alters the associated ground water system. The inflow also may impact nearby surface water regimes, depending upon the interaction of the surface water and ground water systems. The inflow of water into a mine inevitably causes a decline in the ground water levels in the vicinity of the mine. This decline may be restricted to a small area or it may extend a considerable distance, depending upon the hydrogeologic properties of the associated hydrostratigraphic units. A decrease in surface flow will result if the area of ground water level decline intersects a stream or spring that is hydraulically connected to the ground water system. Mine inflow problems historically have been evaluated after-the-fact. Premining hydrologic investigations only recently have been initiated. Brown's (1977)* report on anticipated hydrologic impacts of oil shale mining in the Piceance Basin of Colorado is an example of premining hydrologic planning. Observed difficulties inherent in designing a data collection network and then utilizing the data to predict mine inflow have led us to the preparation of this manuscript. This manuscript reflects a two-phase effort. The first phase describes and develops analytical and numerical techniques for mine inflow prediction. Emphasis in this section is placed on the selection and application of data collection techniques and data analysis procedures. The second phase applies the prediction technique(s) developed during phase 1 to two existing mines. The objective of phase 1 efforts is to evaluate the hydrologic variables associated with surface water systems and ground water systems that can affect the inflow of water into an underground mine. Subsequently these variables are incorporated into predictive techniques for estimating underground mine water inflow and its effect on associated ground water and surface water systems. The specific objectives of phases 1 and 2 are to: 1) Determine the hydrologic parameters required for predicting underground mine water in- flow and the potential effects of the inflow on surface and ground water systems. The hydrologic parameters represent both premining and past initiation conditions. 2) Outline methods for obtaining the necessary input data for the predictive techniques in pre- mining and post initiation periods. 3) Present existing numerical techniques for predicting underground mine water inflow including assumptions Ad required input. 4) Evaluate existing numerical techniques previously tabulated based on their applicability to mine water inflow prediction. 5) Attempt to develop analytical techniques for predicting underground mine water inflow. 6) Develop, through modification of existing models, a general mathematical model as a numerical technique for predicting underground mine water inflow and the resultant effects on surface and ground water systems. 7) Apply the predictive techniques to two existing mines selected for this purpose. 8) Compare the predicted inflow rates to measured inflow rates at both mines. 9) Compare predicted impact on district hydrology to observed impacts where possible. 10) Evaluate the predictive technique.

Jan 1, 1986