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By R. P. Tavares, M. Isac, R. I. L. Guthrie

"The configuration of the metal delivery system used in twin-roll casting can play a significant role in determining both internal and surface quality of the strips produced by this process.In the present work, a mathematical model for coupled turbulent fluid flow, heat transfer and solidification was developed and used in the study of metal delivery systems for twin-roll casting of low-carbon steels. This study was applied to a pilot twin-roll caster being investigated in Canada, with a roll radius of 0.3 m, producing strips of thicknesses from 4 to 7 mm, at roll speeds ranging between 4 and 12 m/min. Various configurations of metal delivery systems were simulated and their effects on the advance of the interdendritic solidification front and levels of turbulence and temperature within the caster investigated. According to the predictions of the model, an extended nozzle with inlet flow towards the surface of the rolls seems to be the best alternative for feeding liquid steel into twin-roll casters.A full scale water model of a pilot twin-roll caster was also built and used to qualitatively validate the predictions of the mathematical model.Various techniques for fluid flowvisualization were applied. In general, the predictions of the mathematical model were in reasonable agreement with the experiments with the water model."

Jan 1, 1999

By M. Frenzel, D. Ebert, K. Bachmann, A. D. Renno, R. Möckel, J. Gutzmer, J. Krause

With the increasing demand for metalliferous and mineral raw materials and the consequent depletion of the global natural resource base, the possible utilization of secondary raw material sources is receiving more and more attention. In the present study, we present results from a detailed vanadium deportment study of three basic oxygen furnace slag (BOS) samples known to containing elevated bulk concentrations of vanadium. Complementary analytical methods that were used to quantify the abundance and composition of V-containing phases include SEM-based automated mineralogy, X-ray fluorescence analysis, and X-ray powder diffraction as well as electron probe microanalysis. The vanadium deportment was quantified using Monte-Carlo simulations of the data obtained from automated mineralogy and electron microprobe analysis. The total V concentrations identified by XRF are between 1.7 and 2.2 wt.% V. The most important hosts of vanadium are larnite-, brownmillerite- and portlandite-solid solutions. In two samples Ca carbonates also significantly contribute to the V deportment, while wuestite, lime, and native iron do not contribute significantly to the vanadium deportment. A thorough consistency check identifies considerable uncertainties in the density of the V-bearing phases as the most likely reason to explain remaining discrepancies between measured and calculated V values.

Oct 14, 2023

By Melinda Channel, Alan W. Weimer, Jonathan R. Scheffe

"There is a need to develop cost effective renewable energy processes for the production of clean fuels that can simultaneously mitigate the U.S. energy/global environmental security issue. If H2 can be obtained cost effectively via the splitting of water using concentrated sunlight, then it’s possible to operate a fuel cell using renewable H2 or to synthesize other fuels based on renewable H2 for generating electricity or for running internal combustion engines. For example, H2 can be reacted with sequestered carbon dioxide (CO2) to produce syngas (H2/CO) via reverse water gas shift (RWGS), i.e. H2 + CO2 ? CO + H2O. The syngas can be catalytically methanated to form synthetic natural gas (SNG) or it can be catalytically reformed to produce liquid fuels such as “green” gasoline via the MTG process (syngas? methanol ? gasoline). The desert SW United States is home to world class sunlight. Fundamental ferrite thermodynamics will be reviewed, an example of the beneficial effect of ferrite surface area on reaction kinetics will shown, and process economics will be investigated for a 100,000 kg/day facility to produce H2.Introduction and Background ThermodynamicsSpinel ferrites of the form MxFe3-xO4, where M generally represents Ni, Zn, Co, Mn, or other transition metals, have been shown to be capable of splitting water according to the two step cycle shown below:MxFe3-xO4 + solar thermal energy ? xMO + (3-x)FeO + 0.5 O2 (1)xMO + (3-x)FeO + H2O ? MxFe3-xO4 + 2H2 (2)This is an inherently clean and sustainable process, as the only net inputs are solar energy and water, and the net outputs are hydrogen and oxygen. Ferrite water splitting cycles are advantages compared to other high temperature water splitting cycles, such as the ZnO/Zn and Mn2O3/MnO cycles, because thermal reduction has been demonstrated at significantly lower temperatures[1, 2]. As a result, overall efficiency may be increased due to the fact that radiation losses increase as the fourth power of temperature, and material considerations for reactor design should be more flexible. Additionally, cycle repeatability has been demonstrated with relative ease as compared to other thermochemical cycles [3-6]. However, the amount of hydrogen per mole that is generated is significantly less than for other high temperature thermochemical cycles due to the fact that only Fe3+ is capable of being reduced at the temperatures of interest[6]. In addition, water oxidation kinetics appear to be very slow, likely due to diffusion or surface area limitations created by near surface oxidation [6, 7]."

Jan 1, 2010

By D. Harrington

Ventilation of underground workings consists in establishment of such control of air currents that the underground workers may work in safety, with maximum comfort and efficiency, and without impairment of health; and requires that the mine openings may be made subject to such control of air flow as to remove from these workings at ordinary times harmful gases and dusts, and that, at time of emergency, such as fire or explosion, there may be maintained as much or as little air flow as may be desired covering portions of the mine or the mine in its entirety. Control of air flow is the keystone of any ventilating structure, and this very essential feature is obtainable only by the installation of mechanically-operated fans, together with other;. ventilating devices such as doors, over-casts, regulators, etc. Every mine, large or small, coal or metal, should from the outset be equipped with a fan. While much has been written about natural ventilation and many claims made that in specific mines there is sufficient natural air flow, there are few if any mines, coal or metal, where natural ventilation supplies anything like adequately safe or healthful conditions for underground workers even at ordinary times; and at a time of mine fire or explosion, mines depending upon natural ventilation are practically helpless, and certainly are decidedly dangerous to those unfortunates forced to be in them. While ventilation has practically always been deemed an integral part of coal mining, metal mines have rarely paid much, if any, attention to air circulation until forced to do so by occurrence of some untoward condition or accident. Yet metal mines actually have as great need of efficient circulation of air as have coal mines. The coal mine must remove the dangerous explosive gas, methane, also fumes from explosives and, in occasional places, other gases such as carbon dioxide or nitrogen.

Jan 1, 1933

By L. C. Armstrong

Twice in the past two years and in two widely separated localities—ane near Williamsville, Mo., and the other in the Allard Lake district of Quebec— the Contract Drilling Division of the Longyear Company has experienced perplexing difficulties in drilling granitoid igneous rocks. The rocks in both places are characterized by appreciable amounts of feldspar intimately inter-grown with quartz particles of microscopic dimensions. The disappointingly slow rate of advance and the exceptionally high bortz loss, obtaining during the life of these two projects, are ascribed largely to the resistance to abrasion of these two minerals when arranged in this intergrown pattern. Whether all intergrowth-bearing igneous rocks would offer extraordinary resistance to bit progress cannot be stated with assurance on the strength of these experiences alone. More data on diamond drilling this sort of material in a number of localities are needed before any generalizations can be made safely. It should be pointed out that the lack of time and funds made it impossible for the writer, or other persons interested in correlating the details of bit histories with the kinds of rocks drilled, to be present during the performance of the contracts. However, the records kept in the normal course of drilling are deemed complete enough to be reportworthy. This information is presented here together with the findings made in studies of the rocks involved, with the thought that it may attract the comments of others who may have encountered similar refractory drilling conditions. A pooling of knowledge on the problem would be of material aid in devising more efficient cutting tools and procedures for penetrating this kind of rock. Drilling Blastholes in the Williamsville Granite Due to the unsatisfactory perform- ance of pneumatic percussion drills in boring blastholes on a construction contract near Williamsville, Mo., tests were made with a diamond drill. The poor progress achieved and the excessive loss of diamonds, however, soon led to the abandonment of these trials. The diamonds, in all of the bits tried, lost their cutting edges and acquired a polish after being in service for distances ranging from a few inches to a few feet. Coring bits containing surface-set, whole-stone bortz in a tungsten powder matrix were employed. The cost assignable to bortz wear averaged $7.17 per foot. This item of expense is several times the amount normally expended in diamond drilling in firm, unfractured, granitic rocks. Diamond-impregnated, tungsten carbide bits also were used without success. The diamond wear was at the rate of $0.98 per foot, but the diamond chips became too polished for further duty in less than a foot of drilling. The Williamsville granite is pink and medium-grained. In the hand specimen, it has a fresh, sound appearance; it is seemingly composed almost entirely of interlocking grains of feldspar and quartz and has no discernible features suggesting a greater durability than is usually found in rocks with this composition and texture. Microscopic examination, however, reveals a predominance of feldspar grains with more or less regularly distributed and sharply angular masses of quartz (see Fig 1). The pattern assumed by the two minerals in this mode of occurrence is referred to as a "graphic intergrowth," as it bears some resemblance to the cuneiform characters of ancient Persian writing. The other principal constituents are quartz-free feldspars (microcline and orthoclase) and quartz. All of these constituents are in nearly equidimensional grains, ranging up to about 3 mm in length. There are also

Jan 1, 1950

By George W. Mitchell

THE property of Eureka Corp. Ltd. is located in the approximate geographic center of Nevada, 2 miles from Eureka, the county seat. The great sources of power, the Colorado, Snake, and Salmon Rivers and the rivers of northern California, are 300 to 500 miles distant, and no lines serve areas closer than 150 miles. Fuel for diesel and steam generation is available in Utah, 300 to 400 miles to the east. Eureka's railhead is 80 miles north where two trunk lines cross the county. A spur line serves Ely, 77 miles east. Good highways connect Eureka to the railheads. Activity in the Eureka mining district began in the early 1870's. The oxidized high grade lead-sil-ver-gold ore terminated against the footwall of the Ruby Hill fault, and in 1890 the main operations ceased. In 1938 Eureka Corp. Ltd. discovered ore in the hanging wall of the fault by diamond drilling. The history of Eureka in the late 1800's indicates that there was some water at 600 to 800 ft in the old workings, probably accumulations above the water table which did not seriously interfere with mining operations. Both the Locan and Richmond shafts were sunk to a level below the table, but apparently the only serious difficulty with water occurred in the Locan. The steam pump used when the last work was done on the 1200 level in 1923, many years after exhaustion of the main orebodies, is still installed on the Locan 900 level. The capacity was about 500 gpm, lifting 750 ft to the 100 level, which connected with the surface. In addition to this. bailers were used to keep the 1200 level free of water. It is said that pumping in 1923 lowered the water in the Holly shaft, about a mile and a half away, but this seems doubtful. The pumping was of short duration because no ore was found. When work at the new Fad shaft was started in 1941 Eureka Corp. Ltd. engineers were fully aware of the probability of encountering water in large volume. Their primary exploration and development had to be carried on at the 2250 level. The first water was encountered at 300 ft. This was undoubtedly surface drainage in the bedding of the Pogonip limestone and was less than 100 gpm. The fractured, loose Hamburg dolomite at the water table was not well cemented, and relatively little water, 300 gpm, percolated through it with difficulty. At 1350 ft well-cemented dolomite containing some open fractures was encountered. These fractures produced the first water of consequence, 750 gpm. At 1700 ft the volume was 1000 gpm increasing to the maximum during shaft sinking, 1600 gpm, at the 2000 level. Secret Canyon shale, a dry formation, was entered at 2100 ft, where a concrete water ring was placed to catch all of the water. The volume decreased rapidly to a constant flow of 1200 gpm. Below 2100 ft the shaft and stations remained in the shale and water was not a problem. Several faults of moderate displacement, including the reverse Martin fault, had been intersected during the traversing of 1000 ft of wet Hamburg, but no undue quantities of water were encountered. Observations in the diamond drill holes in the ore zone area showed a rapid lowering of the water table. The shaft was flooded when it left the dry shale and entered the water-bearing Eldorado dolomite on the 2250 level, crossing a fissure which paralleled the Martin fault. High pressure water doubled the volume then being pumped. Pipe failure through a water door bulkhead was a contributing factor. Immediately following this flooding in March 1948 preparations were made to recover the shaft as rapidly as possible by increasing power and pump capacities as needed. Measurements before flooding indicated the water could be lowered at a fast rate. However, the water table did not recede as rapidly as expected and volumes required to lower the water in the shaft were higher. Obviously the size of the main water channel on the 2250 level was increasing because of erosion, allowing greater volumes to enter the workings and draining beyond the cone originally being drained during shaft sinking. Eroded material was being deposited in the shaft below the 2250 level in serious proportions. In December 1948 a second flooding of the Fad shaft was allowed for the purpose of reassessing existing conditions and studying alternate methods of attack. The detailed geology of the Eureka mining district, see Fig. 1, has been described during the past 75 years by many geologists.' Only the general features and those which seem to affect the drainage problem will be discussed. The old ore zone, mined between 1870 and 1890, is located in a wedge-shaped block of Eldorado dolomite between the footwall of the Ruby Hill fault and the underlying Prospect Mountain quartzite, see Fig. 2. Production of high grade oxidized lead ore containing high values in gold and silver has been variously estimated at $50 to $90 million. The tonnage mined was probably close to 1,500,000, nearly all of which was found above the water table. The new ore, discovered by diamond drilling in the hanging wall of the Ruby Hill fault, is a flat-

Jan 1, 1954

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By J. L. Carne

ORE control is a matter of planning and supervision based on a foreknowledge of the content and distribution of ore. The Inspiration orebody is predominately a copper-sulphide blanket, overlain by an oxidized zone of copper silicates, copper carbonates, and barren capping. Most of the ore is treated by leaching, and the optimum requirement of this method is the goal toward which the mining schedule is aimed. Both open-pit operations and underground mining are used. The problems of ore control at Inspiration are: I—maintenance of metallurgical balance, 2—conservation of oxide ore, and 3—mining out of the orebody at an average grade which parallels that of the reserve figure. Prior to the fall of 1926 all of the Inspiration division ores were treated by flotation, but the mixed ores of the Live Oak—Keystone section carried too large a proportion of copper in the form of chrysocolla for satisfactory flotation. For this type of mixed ore the leaching process now in use had been worked out, and the plant was put into operation in the fall of 1926. The process is known as the ferric sulphate leach, and the leaching solvents contain both ferric sulphate and free sulphuric acid. Sulphide copper in the form of chalcocite is leached by the ferric sulphate, and the oxidized copper minerals are dissolved by the sulphuric acid. From the beginning of operations to the present time the plant has treated some 67 thousand tons of ore having an average copper content of 1.142 pct, of which 0.605 pct has been in the form of oxidized copper minerals. The plant is currently being operated at the rate of 11,500 tons per day, 7 days per week. However, this tonnage is produced in 6 days of mine operation. The average current grade is approximately 1 pct copper, of which about 50 pct is present as oxidized or acid-soluble copper. Close control of metallurgical procedures is a vital factor in operation of the plant and is in turn dependent upon careful regulation of feed. An excess of sulphide copper requires the presence of more ferric sulphate in the leaching solvents. On the other hand, an excess of ferric sulphate in the presence of a low sulphide feed seriously reduces the efficiency of electrolytic precipitation in tank house operations. This condition must be avoided at all costs. An additional factor must be considered. It is known that the sulphide content is predominant in the ore which remains in the reserve. To maintain the necessary sulphide-oxide proportion, so that the entire reserve will mine out at both a grade and a sulphide-oxide ratio that will be permissible throughout the remaining life of the property, it is obvious that the oxide reserve must be currently preserved to as great an extent as possible. Fortunately much of the oxide reserve is found in the open-pit operation. This happy circumstance adds flexibility to the procedure and makes it possible to maintain the close regulation necessary to insure best overall results, both mining and metallurgical. The two ore streams from the pit and underground operations join at the coarse-crushing plant, and proper control must be exercised before the ore reaches this point. The open pit is currently providing about 55 pct of the ore and underground operations the balance of production. Both operations are under a single superintendent so that the closely interrelated problems may be more readily controlled. The least flexible of the two producing units is the underground mine, where the mining method is block caving. Although a block of ground may be undercut and developed, characteristics of the material determine its caving action. The ore in the chutes must be drawn. The area currently mined from underground is in a zone predominately sulphide and is generally

Jan 1, 1953

By W. C. Williamson, J. L. Shugert

DUST control is receiving considerable attention today by most companies as is evidenced by the many articles, papers, and books written on the subject. ~~~t of the larger mining companies have given this subject close study and have established standardized methods of control. W. C. WILLIAMSON and J. L. SHUGERT, Members AIME, are Assistant Ventilation Engineers, Anaconda Copper Mining Co., Butte, Mont. San Francisco Meeting, February 1949. TP 2705 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Jan. 17, 1949. Permissible Limits: The question of permissible dust limits has been considered by medical commit-ties and various national conferences. Dr. Lanza, of the Metropolitan Life Insurance Company, Dr. Le-roy U. Gardner, of the Saranac Laboratory for the study of Tuberculosis, and Dr. Pancoast, of the University of Pennsylvania, have been members of these committees. A statement included in reports of some of these committees is as follows: Concentrations to which the dust must be induced in order to be safe have not been absolutely determined. There is evidence that a concentration of more than five million particles per cubic foot of a highly siliceous dust is dangerous. Therefore, it may be stated that it is desirable to avoid counts of more than five million particles of dust containing free silica to a high percent in working places. Table I gives the safe limits adopted by several state industrial codes. Table I. Safe Limits of Dust Concentration Limit Million Particles State Per Cu Ft Cal~fornia Silica (less rhdn 10 pct) 50 Silica (10 pct to 50 pcr) 10 Sllica (more than 50 pct) 5 New York Frcc silica less than 10 pet by weight of rock formation 100 Frec silica more than 10 Pct by weight of rock formation 10 Oregon Silica (over 40 pct) 5 Silica (20 to 40 pct) 10 Silica (10 to 20 pet) 20 Wisconsin 15 million particles, under 10 nicrons in longest dimension per cu ft of air, when the quartz content of the dust is 35 pct, variations in free silica content will make proportional inverse changer in this standard Close attention to dust concentrations has been given to mining operations, since these operations are frequently carried on at considerable distances

Jan 1, 1951

By J. L. Shugert, W. C. Williamson

DUST control is receiving considerable attention today by most companies as is evidenced by the many articles, papers, and books written on the subject. ~~~t of the larger mining companies have given this subject close study and have established standardized methods of control. W. C. WILLIAMSON and J. L. SHUGERT, Members AIME, are Assistant Ventilation Engineers, Anaconda Copper Mining Co., Butte, Mont. San Francisco Meeting, February 1949. TP 2705 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Jan. 17, 1949. Permissible Limits: The question of permissible dust limits has been considered by medical commit-ties and various national conferences. Dr. Lanza, of the Metropolitan Life Insurance Company, Dr. Le-roy U. Gardner, of the Saranac Laboratory for the study of Tuberculosis, and Dr. Pancoast, of the University of Pennsylvania, have been members of these committees. A statement included in reports of some of these committees is as follows: Concentrations to which the dust must be induced in order to be safe have not been absolutely determined. There is evidence that a concentration of more than five million particles per cubic foot of a highly siliceous dust is dangerous. Therefore, it may be stated that it is desirable to avoid counts of more than five million particles of dust containing free silica to a high percent in working places. Table I gives the safe limits adopted by several state industrial codes. Table I. Safe Limits of Dust Concentration Limit Million Particles State Per Cu Ft Cal~fornia Silica (less rhdn 10 pct) 50 Silica (10 pct to 50 pcr) 10 Sllica (more than 50 pct) 5 New York Frcc silica less than 10 pet by weight of rock formation 100 Frec silica more than 10 Pct by weight of rock formation 10 Oregon Silica (over 40 pct) 5 Silica (20 to 40 pct) 10 Silica (10 to 20 pet) 20 Wisconsin 15 million particles, under 10 nicrons in longest dimension per cu ft of air, when the quartz content of the dust is 35 pct, variations in free silica content will make proportional inverse changer in this standard Close attention to dust concentrations has been given to mining operations, since these operations are frequently carried on at considerable distances

Jan 1, 1951

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By Carville E. Sparks, Rollin Farmin

TRIAL installations of rock bolts, of the slit-rod-and-wedge type, were under way at several units of Day Mines, Inc., when Korean hostilities interrupted the already slow deliveries of steel bars to the Coeur d'Alene district. Factory-made bolts had not yet been put on the local market, so the program was halted for lack of supplies. Interest was revived by a visitor's description of wooden roof bolts. These were said to have been used briefly with apparent success in a coal mine, until apprehension voiced by the U. S. Bureau of Mines caused the practice to be suspended. To make wooden bolts for trial in ground support, Day Mines acquired a second-hand doweling machine equipped with two cutting heads, one to turn out the desired round rods of 2-in. diam, the other to turn out 1-in. rods to be used as powder-tamping sticks. This machine was installed in the all-weather sawmill of the Hercules mine unit at Burke, Idaho, where fabrication of the wooden bolts commenced early in 1951. Most of the mining in the Coeur d'Alene district is along steeply dipping veins in shaly quartzite and argillite of Algonkian age. Ground support commonly is required in zones where the rocks have been sheared, brecciated, and hydrothermally altered. Pressure from the sidewalls is more troublesome than weight overhead, but both increase with the size of the mine opening. Caving may come from a progressive sloughing of irregular rock fragments or from an exfoliation and buckling of the layered wall rocks. The disintegration is thought to develop from an initial elastic expansion of the rock toward the newly-created mine opening, followed by the dilation of many tiny partings in the rock by absorption of water. As the partings widen, masses of rock develop weight and become free to fall. The function of rock bolts is to prevent or retard widening of partings in the rock supported. Wooden Bolts, Wedges and Headboards Bolt assembly used by Day Mines consists of a bolt 4 or 6 ft long, two wedges 16 in. long, and a headboard 30 in. long, Fig. 1. All four pieces are made of local red (Douglas) fir, either green or well-soaked in the mill pond before it enters the sawmill. Bolts are fabricated from cants, 2 1/4 in. sq, cut from relatively straight-grained timber with a minimum of knots and trimmed to 4- and 6-ft lengths. The bolt then is turned in the doweling machine from 21/4 in. sq to 2 in. diam round, except for a 4-in. length at one end which is left full square to provide the striking head and the shoulder that holds the headboard in position for wedging. The foot end of the bolt is slit with a thin saw for a length of about 16 in., thereby making a slot to receive the wedge against which the bolt is driven for anchorage at the bottom of the rock hole. A similar slit, 12 in. long, is made in the opposite (head) end of the bolt to receive the second wedge, which crowds the headboard against the ground at the collar of the rock hole and puts the bolt in tension. The second slot is aligned 90" from the plane of the first slot to avoid Longitudinal splitting and is notched out slightly to allow easier insertion of the collar wedge after the bolt has been driven to bottom. To prevent splitting the headboard by spreading action of the head wedge, this slot is oriented at 90" to the grain of the headboard when the pieces are assembled, Fig. 2. The wedges are similar to standard mine wedges, but more slender; they are cut 1 7/8 in. wide and 1 in. thick at the heel and taper out in 16 in. of length. The headboard, or bearing plate, is not necessary for some types of ground but generally is desirable because it helps the bolt to support an area of loose, friable rock and reduces the tendency for the rock at the collar of the hole to split away from the wedged head by distributing the pressure over a wider rock surface. The headboard may be a 24-to 30-in. length of 3-in. plank, 8 to 12 in. wide, but a similar length of rounded sawmill slab serves equally well at 20 pct of the cost. A hole of 2-in. diam is bored or punched through the center of the headboard, either at 90" or at various high angles to its surface. The bolt is inserted to its shoulder through this hole, then driven into the rock hole. Bolts, wedges, and headboards are given a full timber preservative treatment to inhibit rot. Bundles of each are immersed in a warm saturated solution of Osmose salts in water for 48 hr, removed, dripped dry, and stored in a relatively humid underground depot to cure. Most wooden rock bolts used by Day Mines are 4 ft long. Holes to receive them, about 42 in deep and 2 1/8 in. in diam, are drilled into the rock' to be supported, nearly normal to the periphery of the mine opening. The type of drill used is dictated by convenience: stoper, jackleg, or jumbo-mounted drifter. Correct depth of the hole is assured by use of a measuring stick that has been cut to the proper distance from drill chuck to the ground at the collar of the hole when a standard length drill rod is at the bottom. The bolt is seated to the shoulder through the hole in the headboard, the foot-end wedge is placed in its slot, and the assembly is inserted into the rock hole. Then the bolt is driven until it is seated solidly on the wedge against the bottom of the rock hole. Driving may be by hand with a sledge, or

Jan 1, 1954

By S. H. Ash, J. J. Forbes

Probably the newest thing in mining safety, or safety for mines, is the apparent dissatisfaction on the part of the mineral industries, as represented by both management and labor, and the general public of the Nation, with the safety record of the mining industry. This is reflected by published comments and papers in the technical press;l-9 in labor periodicals,10,11 in popular journals and periodicals,l2, 13 and in the daily press; also by discussions of the Safety Committees sponsored by the National Safety Council,9, 14-15 the American Mining Congress,l.l6 the American Institute of Mining and Metallurgical Engineers,2,4 the Mine Inspectors' Institute of Amerthe Coal Mining Institute of America,5 19 the American Standards Association, and others.20.21 Mining safety is a joint problem of management and the individual workman. Safety is not attained and will not be attained solely by laws, decrees, and commands because, if it could be done in this manner, acceptable mining safety would have been achieved long ago by the individual states in our Nation and by the nations abroad. The best safety performance is found at plants where management provides a safe environment for the workers; where supervision is efficient and super- visors, others concerned with promoting safety, and the individual worker know through training, experience, and personal contact the safe and unsafe practices of their field; and where the trained individual worker has become safety-minded by experience, quiet influence, unconscious suggestion, and personal guidance. Despite what may be said to the contrary, progress has been and is being made in mining safety. This is shown in Fig 1 to 4. It is beyond the scope of this paper to give possible reasons for the valleys and peaks in the fatal-injury rates. Some recent factors that appear to be contributing to the improvement in these rates are briefly discussed. Aside from mining equipment, "what's new in mining safety" for some mines has been in effect or was tried by others long ago. A safety pro- gram is a living thing; as such, it must be vitalized and continuously fed with enthusiasm, constructive action, and improvements. Because safety is concerned with human conduct, it is necessary to practice safety at all times; if not, it is soon forgotten, and accidents occur. Mining-safety Records The large employment and production concerned with anthracite contribute much to the mining-safety records of the mineral industry. Approximately 81,000 persons are. employed in this industry having mines concentrated in an area of 480 square miles in northeastern Pennsylvania. Because of the achievement of the anthracite industry in safety during 1948, the other branches of the mineral industry can profit by a study of what has been done to achieve that record. Where one mineral is produced by other large groups having methods of mining that are similar in many respects to the anthracite-mining methods, an achievement in such a large industry is important. Salient factors concerned with safetye in the anthracite industry are: no radical change has occurred in the mining-safety regulations pertaining to anthracite mines. Because anthracite mining is a branch of the coal-mining industry,

Jan 1, 1950

By Wallace R. Horn

"MR. CHAIRMAN AND GENTLEMEN :I begin this talk by making a couple of comments which are of the nature of disclaimers.Firstly, and as the Chairman has otherwise indicated, I was neither ""born into"" nor have I spent more than a few years of my professional life in the mining industry. I mention this because many might, and perhaps rightly, consider that an added qualification in presenting a report of what the industry has accomplished. Indeed, I am not here to blindly defend all of its actions and inactions in respect to environmental control over the years. Nor am I here to seek absolution of the industry's past transgressions - if indeed that is the word. Mistakes might be a better one. I concluded some time ago that the effects of those mistakes are too difficult to assess against the sea of errors of our total society. I should mention too that I am not an environmental engineer or specialist - only a portion of my interest is involved with this field. Perhaps this permits a perspective which allows a more accurate and objective view.One further comment here. My job today, as I understand it, is to generally report upon the nature and extent of the response of the metals mining industry to control of the environment. In the case of an industry operating at more than a hundred locations and for various reasons at various stages of progress, it is not an easy job. On the other hand, I wanted to avoid the use of generalities because, especially within the context of environmental or ecological opinion and of the practice of condemnation of industrial segments, I think we've had about enough of generalities until we know far more than we do.Finally, I am quite aware that a statement of future plans to achieve environmental control does not impress or satisfy as many people as it may have done some years ago. Public opinion demands to know what is currently underway, today."

Jan 1, 1972

By FREDERICW K. BRADLE

I HAVE been puzzled by two lines of thought'; one emanating from Washington, D. C., to the effect that we must all cheer up, that in a very short time, measured in terms of months, prices would begin to recover, business would become normal, and prosperity would continue on its way just as before the interruption of a year ago; and a second, coming from certain industrialists and economists, expressing fears that all commodity prices were in for a long downward trend, which would cause much unemployment, eventually drag down wages and so restrict purchasing power as to further aggravate the situation. One basis for the fear of a downward trend in commodity prices' is that ifter great wars a depression could last for something like 40 years-the downward trend to such a depression after our recent World War having been postponed for some reason or other; and the basis for another fear was that the production of gold was becoming less. I had 'also been trying for some time past to refresh my recollection as to how my ow11 particular lead interest climbed out of the depression of 1893; and as all the metals have had, and still have. relative values and prices, I can best express what I have in mind by talking principally about lead.

Jan 1, 1930

By L A. Martinez Tipe, J McKibben, S Barry

"Metal prices are central to a mineral investment decision and have a profound impact on the financial performance of companies in the mining industry. At its most basic level, metal prices are influenced by the disequilibrium between international metal volume supply and demand. That is, the commodity price is a regulator that equates international metal supply and demand where inventories act as a buffer between the emergence of this disequilibrium and its dissolution through a price change.From this viewpoint, mining companies are required to forecast metal and bulk commodity prices when evaluating minerals projects through discounted cash flow (DCF) analysis. While supply-demand models are instructive, many companies use historical prices, spot prices, futures/forward pricing and consensus opinion (or a combination of all of these) to develop price forecasts.This paper discusses the definition, fundamentals and differences between futures/forward and consensus pricing and the appropriateness of each in a DCF model for mine project economic evaluation.Our results demonstrate that great care is required when using either futures/forward or consensus pricing and that the decision to use one rather than the other requires consideration of other important elements in the DCF model (ie the risk-adjusted discount rate).CITATION:Martinez Tipe, L A, McKibben, J and Barry, S, 2016. Metal prices and the differences between forward and consensus pricing – which one is better for use in a discounted cash flow model?, in Proceedings Project Evaluation 2016, pp 149–158 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

By Y. X. Yang

Bottom ash is a solid waste stream in a Waste-to-Energy (WTE) processing plant during the combustion of the municipal solid waste (MSW). In the Netherlands, bottom ash contains typically about 18% metals originated from the MSW, and current industrial practice could recover about 10% metals through magnetic and eddy current separators. The remaining 8% as the finer particles is difficult to recover through current physical separation technology. Among various alternatives, vitrification is one of the promising options to immobilize the ash and recover the remaining metals. During vitrification a homogeneous glassy slag and a Fe-Cu based alloy were obtained. In the cur-rent study, the metal alloy recovered from the bottom ash originated from a Dutch Waste-to-Energy plant contains about 82 wt% Fe, 12 wt% Cu and 6 wt% other alloying elements and impurities such as C, S, P, Ni etc. The generated alloy does not have direct commercial applications, and separation of copper from iron-based alloy is critical to obtain the market value of both copper and ferrous constituents. To find a suitable processing route of this metal alloy, various alternative approaches have been considered for an efficient removal of copper from the alloy, and the focus in the present study was on the sulphide treatment at elevated temperatures of 1400 ~ 1500oC. The tested sulphide systems include pure FeS, FeS containing sulphide mixture (such as Na2S and Al2S3), pyrite (FeS2) mineral, ZnS containing sludge as an industrial waste. It was found that the carbon saturation of the alloy is proved to be an essential condition for a proper phase separation. The results showed that FeS mixed with Na2S, ZnS containing sludge and pyrite are effective sulphide systems for copper removal with a removal efficiency of 92% by using pyrite and 95% by zinc sludge at a S/Cu ratio of around 7 times of the stoichiometry. The copper content drops from 12 wt% to about 2 wt%. However, it is still difficult to remove copper completely from the alloy in a single step, and a multi-step operation is needed. The present research is expected to contribute both to the utilization of the metal recovered from bottom ash, and to the general recycling of steel scrap where copper removal from the steel scrap is a critical issue.

Jan 1, 2014

By R. Boom, C. F. Balentina, J. H. L. Voncken

During vitrification of bottom ash generated from the combustion of municipal solid waste (MSW), a Fe-Cu based alloy phase with great value was formed. In order to find proper commercial use in metallurgical industry, it is necessary to separate Cu and Fe. The present study focuses on the sulfide treatment at temperatures of 1400 – 1500oC, in particular with FeS containing minerals. Various sulfide compounds such as Na2S, Al2S3 and Ni3S2 were added to promote the copper removal. Pyrite and zinc sludge (containing ZnS) as waste streams were studied with specific interests. The results showed that both zinc sludge and pyrite are effective for copper removal with the highest removal efficiency of 95% by using pyrite. The copper content drops from 12 wt% to about 2 wt%. However, it is still difficult to remove copper completely from the alloy in one step, and a multi-step operation is needed.

Jan 1, 2008

By A. Janwong

Ion exchange technology is becoming more prevalent for metal recovery from low concentration solutions. The products of ion exchange are typically a metal free stream (product stream) and a concentrated metal stream (byproduct). The metals in the concentrate stream are usually precipitated and returned to the process or discarded. The unique features of emew® technology enable metals to be recovered directly from the concentrated stream. The advantage of the IX-emew® combination is the ability to have two product streams: 1) clean metal free solution and 2) metal product for sale. In addition to recovering metals from solution, it is also possible to separate metals from each other and recover more than one metal product. In this paper, examples of IX-emew® combinations are detailed including the recovery of copper, nickel and cobalt from an industrial waste stream, recovery of nickel from a spent electroless nickel plating solution and recovery of cobalt from an hydroxide cake.

Feb 23, 2014