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ALABAMA. Aldrich, Aldrich, W. F. Anniston, Noble, A. E Bessemer, Abbott, C. E., Fer¬gusson, V. Birmingham, Aldrich, T. H., Jr., Bowron, C. E., Bowron, J., Collord, G. L, Crockard, F. H., Davis, F. H, Dimmick, F. D., Hillman, Ludwig, C. P., McCarthy, M. E., McCormack, G B., McCrery, Ramsay, E., Shook, A. M. Car¬bon Hill, Pill, J. R. Coal City, Lydecker, I. S. Gannts Quarry, Hiller, E. Helena, Squire, J. Holt, Shook, J. \V. Huntsmlle, Moss. Lehigh, Toulmin. Oxmoor, Green¬amyer, A. G. Rock Run, Garvin, J. M. Sheffield, Miller, D. I. Woodward, Lodge, J., Woodward, A. H. ALASKA. Ellannar, Stevens, B. Fairbanks, Hamilton, A. L, Robe, L S. Hadley, John¬son, P. Jualin, Hoggatt, W. B. Juneau, Denby, Thane, B. L., Whipple, J. R. Katalla, Hampton, W. H. Ketchihan, Barber, J. C., Geiger, A. \V., Parker, J. L. Niblack, Finck, W. N., Niedling, B. B. Nome, Garfield, C. D. Tin Oity, Fay, A. H. Treadwell City, Kinzie, Stow, H. P. Valdez, Esterly, G. M., Hazelet.

Jan 1, 1910

By R. V. Peche

Since 1952, Bond’s equation has been used as a standard to predict the specific energy of grinding for a given duty. The Bond’s equation considers that the exponent is constant for all conditions. Hukki, in 1961, pointed out that the exponent is a function of the particle size, and concluded that the Bond’s equation applies successfully for an intermediate range of sizes; out of this range the Bond’s equation leads to high error as demonstrated in this paper. More recently, in 2004, S. Morrell presented a new formula in which the exponents vary with the feed and product size. This paper is based on Hukki’s approach and presents a new general formula, in which the exponent varies only with the feed particle size. This new equation has been proven in 82 observable cases in Peru and other countries around the world, involving ball mills and vertimills in different configurations over a wide range of feed and product sizes, work indices and mill diameters; in most cases, the predicted value has an error less than 10% with regards to the observed value and no need for any correction factor. Keywords: Grinding, specific energy, particle size, work index, variable exponent

Jan 1, 2020

By Mike Muchard

A design phase load test program was performed at the Broadway Viaduct bridge site to evaluate load carrying capacity and constructability of post grouted drilled shafts. The program included construction, instrumentation, post grouting and load testing of three drilled shafts. Two shafts included sleeve port and plate base grouting apparatus and were post grouted. The third shaft was un-grouted and served as a control shaft for comparison of the un-grouted end bearing. The shafts were 5 feet in diameter with the control shaft at 55 feet in length and post grouted shafts at 55 and 66 feet. SPT borings at the centerline of each test shaft defined geotechnical conditions at each shaft. Test shafts were instrumented with strain gages and concrete integrity was evaluated with crosshole sonic logging. Instrumentation and data acquisition was used during the post grouting operation to fully capture the shaft behavior. Finally, Statnamic load testing was performed on all three shafts to their ultimate capacity. The fine to coarse sandy soils in which the shafts were tipped in were well suited to improvement from post grouting. Grout pressures of 660 psi were achieved with 0.08 inch or less of upward shaft movement. Subsequent Statnamic load tests showed 183 to 192 percent greater end bearing in the post grouted shafts as compared to the un-grouted shaft. The overall result showed a potential savings of approximately 9 feet to 35 feet of drilled shaft length in addition to the increased quality assurance from post grouting.

Jan 1, 2009

By Jannie Maree, Dieks Theron, Ryneth Nengovhela, Patrick Hlabela

In extracting base metals from ore bodies, sulphide as well as oxide, the main environmental concern is sulphur, whether in the form of sulphurous gases or sulphate in liquid form. Generally sulphur is in over supply in developed countries mainly because of strict legislation to capture it. In developing countries on the other hand there is usually a shortage of sulphur because of less strict environmental legislation. Africa falls in this second category and the African continent imports substantial amounts of sulphur despite the fact that large quantities of sulphur is emitted by smelters, thermal power stations, coal mines, fertilizer manufacturers and others. Other than making acid from gaseous emissions, existing methods to capture sulphur from gaseous and liquid effluents are uneconomical and it ends up in the environment either as ammonium sulphate (from gaseous effluents) or as gypsum (from liquid effluents). The CSIR?s Environmentek and Pyroserv has developed a process whereby gaseous sulphur emissions and sulphates in liquid effluents can be economically captured and converted into elemental sulphur and metal carbonates. The elemental sulphur is of high purity as it is precipitated from a gas stream and is readily marketable. The metal carbonates (barium or calcium) can be recycled back to the process or marketed. The only harmful byproduct of the process is carbon dioxide. Should natural gas or methane be available the carbon dioxide could be reduced by about one third. Should solar, wind or hydro energy be available to generate hydrogen the carbon dioxide emissions would be zero. Employing this technology could make Africa independent of sulphur imports and furthermore reduce the impact of acid rain and seepage of sulfate from gypsum and other waste dumps into the environment.

Jan 1, 2005

By William E. Bruce

Percussive drill calibration experiments performed by the Bureau of Mines on two drills are described. The use of a linear velocity transducer to determine piston velocity, blow frequency, and piston stroke length is discussed. Data are presented on the relationships between maximum work rate and feed air pressure. Penetration rates in nine rock types were measured for the same two drills. Physical properties of the nine rock types are presented in tabular and graphical form to show relationships between penetration rates and physical properties. Finally, conclusions are drawn as to the potential use of these data to predict percussive drilling rates.

Jan 1, 1970

By E. H. Cho

A current environmental concern is the cleanup and restoration of soil contaminated with low-level radionuclides at DOE sites including the Nevada Test Site. At the present, there is no applicable process to economically and effectively deal with the remediation of contaminated soils. This situation requires development of a selective flotation process to separate the radionuclides from contaminated soil efficiently and economically. In this study, a minus 200 mesh size fraction of pure cerium dioxide was used as a surrogate for plutonium oxide which is the major contaminant at the Nevada Test Site. A Hallimond flotation cell was utilized to examine the flotation behavior of cerium dioxide with oleic acid under various conditions of pH and collector concentration Mechanical cell tests were conducted to float the surrogate from a mixture with uncontaminated Nevada Test Site soil in the absence and presence of calcium carbonate as a carrier. Two different sizes of the carrier (minus 150 mesh and 65-150 mesh) were tested. It was found through the Hallimond cell flotation tests that the cerium dioxide could be floated well at pH's 4 and 9 with 10-3 molar oleic acid. It was also found that the addition of the carrier drastically enhanced the flotation recovery of cerium dioxide. The flotation recovery of the surrogate increased from 34 % to 65 % by adding only 0.5 % calcium carbonate (minus 150 mesh) in the feed and also increased to 95 % by adding 10 % of this carrier to the feed. The finer size fraction of calcium carbonate (minus 150 mesh) was more effective in recovering cerium dioxide than the coarser size fraction (65-150 mesh) The flotation method utilizing a carrier of calcium carbonate seems to provide a high potential to be used for the remediation of soil contaminated with radionuclides at various Department of Energy's nuclear weapon sites.

Jan 1, 1996

By T. K. Mukherjee

The importance of the processing of secondary resources such as solid industrial wastes and liquid effuents as well as byroducts for the recovery of the associated metal values hardly needs any elaboration. Such processing contributes in no small measure.towards the augmentation of the availability and the supply position of resources. As an added benefit, it leads to a reduction in environmental pollution caused by the undesirable waste materials. As some typical examples of secondary resources in the Indian scenario, mention' can be made of spent catalysts, metal/alloy scrap, ash, consumed batteries, flue dust, slag, sludge and slime, pickling solutions, spent acids; plating rinse solutions, spent etchants, waste water from synthetic fibre manufacture and from tanneries etc. These wastes constitute a source of a variety of non ferrous metals. In the processing of secondary resources like these, the techniques of hydrometallurgy such as solvent extraction play a pivotal role. This paper outlines the efforts made to use solvent extraction to recover (1) base metals such as copper, zinc, nickel and cobalt from a number of secondary resources like cake, and sludge(2) refractory metals like molybdenum and vanadium from sulphidic, byproducts and spent catalyst and (3:) nuclear metal like uranium from phosphoric acid. In the case of base metals, process flowsheet involving the use of acidic extractants such as DEHPA and PC88A have been optimised. The problem of iron removal from the leach liquor, "which influences the metal recovery process significantly, has drawn a special emphasis. hi the case of refractory metals, basic extractants .like Alamine 336 has been pressed into service to produce kilogramme. quantities of pure salts of molybdenum and vanadium, Finally, the paper gives a brief account of plant experience in recovering uranium from phosphoric acid, using 'a solvent system comosed of DEHPA and TOPO.

Jan 1, 1996

By T. W. Sullivan

"This paper explores the history of AC static variable frequency control drive systems on open cast mining machinery, in particular electric mining shovels and electric walking draglines, over the past ten years. It critiques the peformance of these drive systems from a productivity and operating cost standpoint and evaluates the over-all success of the original design concepts as they were envisaged in the 1970s. The paper details the state-of-the-art as it exists today as well as design improvements contemplated for the immediate future.IntroductionConsidering the fact that the very first successful application of an alternating current electrical distribution system occurred at the Gold King gold mine near Ames, Colorado, U.S.A., in the year 1891, it seems somewhat ironic that almost 100 years later, AC excavator drive systems are just now becoming commonplace in the mining industry.If one were to read the biography of Lucien L. Nunn, the first person to successfully install a commercially viable alternating current power distribution system, one would quickly visualize a parallelism between his fight to achieve acceptance of his new lower cost technology and that of those commissioned with the belief that AC static variable frequency control drive systems are not only a better performing, lower cost drive system for mining excavators today, but moreover, the drive systems of the future in our industry.This paper will deal with the evolution and incorporation of AC static drive systems on open cast mining excavators. It will trace the history of the development of these drive systems in the mining industry, the current state-of-the-art concerning these systems, and what the future holds for these excavators in our industry."

Jan 1, 1990

By W. C. Phalen

INTRODUCTION NOBODY will dispute the fact that the conservation in every legitimate manner of our valuable high-grade phosphate-rock deposits is a present-day problem of importance. The table and curve, given herewith, show. that during the past 10 years the exportation of high-grade rock has averaged very close to half the total output of the entire country. The bulk of this exportation is from Florida, for obvious reasons. It is plain that the deposits in this State, more particularly, are being wastefully depleted under a system of selecting the cream of the product for exportation to Europe, leaving the comparatively low-grade rock, running from 65 to 70 per cent., and under, in bone phosphate of lime, for our own n fertilizer manufacturers to work up when all the best rock is gone. In this paper the writer has endeavored to make a point of describing in some detail the important methods of production and conservation in Tennessee. These ought to prove of educational value to our American agriculturists and fertilizer manufacturers, and should result in a demand for the highest-grade rock, both for direct application to the soil and for use in making acid phosphate. It is certainly evident that the European

Jan 10, 1916

By E. M. Sacris

The semi-autogenous grinding mill is the bottleneck whenever treating hard and finely fragmented ore at Benguet Corporation's Kaline Concentrator. Low milling rates result from very high and fluctuating mill power draw. Attempts to use a mill horsepower control loop to regulate feed rate were unsuccessful because of operational upsets caused by fluctuations in feed rate on downstream operations. The objective of the computer control was to maximize tonnage throughput by stabilizing the power draw at a high but safe level without upsetting subsequent operations. The semi-autogenous mill is operated in closed circuit with a stationary screen. Instrumentation consists of a feed weightometer, a power transmitter to measure mill load, a headwater flowmeter and an ultrasonic level sensor to measure the mill discharge sump level. A 64 k-byte FOX-3 Foxboro computer processes field data from various sensors and manipulates the input control variables. These variables are the fresh feed rate and the headwater flowrate. Feed rate is controlled through a supervisor-type computer program aimed at maximizing and stabilizing power draw. The control scheme was according to how the operators responded during an operator-directed automatic control mode. Mill dilution is similarly controlled. Results achieved reflect a 3% increase in tonnage milled because of the reduction in control variability that ensued.

Jan 1, 1987

By William J. Douglas

Critical path modeling (CPM) has been used widely as a tool to manage large-scale projects having many activities which are conducted in series or parallel. The construction industry is a regular user of critical path modeling as a planning tool. In this context, the method is used to provide an orderly schedule for all activities to forecast delays, slippages, and problem areas and to identify critical activities where resources should be directed. It also identifies others which have slack time, and provides information on the degrees of freedom and flexibility available in completing the project. The purpose of applying critical path modeling to an underground mining system is not primarily for scheduling, but as an analysis tool. In observing underground mining systems, it Is clear that repeated sequences of activities must take place for mining to proceed efficiently. Some of these activities must be sequential, while others be carried out in parallel. The fundamental difference between critical path modeling applied to mining versus a large-scale construction project are: (1) The activities in a mining system are cyclic in nature. (2) The location of the activity varies, and the entire process is not thought of as co-located, as is the case in a construction project. Another important aspect of the network development for mining systems is the degree of detail that is desired. One can be concerned with very macroscopic networks which represent, for example, an entire mine life cycle. Networks representing this problem would be more closely analogous to those of a large construction project. On the other hand, one might be interested in evaluating alternative ways of developing a property or of mining in a given section. More detailed networks representing the section activities are necessary to evaluate system parameters and their impact on long-term production.

Jan 1, 1978

By Forsén. O.

In this ?paper practical solidification of anodes containing nickel, arsenic, antimony and bismuth is compared with the equilibrium? diagrams and kinetic ?factors. In order to clarify the distribution of the impurity elements in the anode the structures were studied after the solidification and also after the electrolysis of anodes. This examination was carried out with optical microscope, scanning electron microscope and electrochemical measurements. All the nickel is remaining metallic in the copper matrix during the solidification and therefore nickel is dissolved in the electrolysis at all oxygen contents, when the nickel content is 0.25 wt-% or less in the anode. The practical .solidification mechanism of the anodes, when the nickel content exceeds 0.25 wt-%, proceeds by the crystallization of nickel oxides. In oxygen containing copper anodes nickel oxide is verified to crystallize together with antimony in certain conditions and to form a micaceous compound (3Cu2O?4NiO?Sb2O5), which may build up a very harmful floating slime in the electrolyte. Arsenic is strongly enriched along the phase boundaries in oxygen free copper anodes. An intermetallic compound Cu3As is formed when the amount of arsenic in the anode. exceeds 0.7 wt-%. This compound is more noble than the copper matrix. This means 'that a selective dissolution of copper-arsenic compounds takes place. The copper matrix dissolves more strongly than the grain boundaries. The solidification structures of oxygen free bismuth containing copper anodes show clearly the influence of bismuth. Thin bismuth films which are less .noble than the metallic copper matrix, are formed during the solid1ficatton process. As a result of this the?grain boundaries are dissolved faster than the copper matrix.

Jan 1, 1987

By P. R. La Pointe, H. C. Ganow

The production of both over- and undersized coal fragments is a constant problem to open pit operators in the Powder River Basin. Operators strive to achieve production blasts that economically minimize the weight percentage of over- and undersize coal. However, natural fractures influence blast fragmentation, and can impose limits on the final fragment size distribution. ARCO&apos;s Anaconda Minerals Company studied the limits placed by natural fracturing as part of a wider geotechnical study of the Wyodak Coal in the eastern Powder River Basin, Campbell County, Wyoming. Measurements of fracture orientation, size and spacing were made at ARCO&apos;s Black Thunder Mine. These data were used to form statistical distributions for natural fracture size, spacing and orientation. An algorithm was created to select fracture properties at random from these distributions and to form hypothetical blocks. The blocks were then numerically "sieved" to derive pos- sible particle size distribution curves. These curves are compared to fragmentation results from samples taken from three test blasts. All test blasts yield very similar lognormal fragment size distributions. The median particle size varies from 1.25 to 2.0 in.(3.2 to 5.1 cm), the mean from 4.0 to 8.5 in. (10.2 to 21.6 cm). The standard deviation was about 1.6 in. (4.2 cm). Coal fines (fragments < 0.25 in. or 0.64 cm) are about 15%. Calculated fragment distributions for the natural fractures predict the median size well but underestimate the variance. Algorithm modifications to account for processes known to occur in the comminution zone near the blast hole significantly improve the match in the fine coal size range. The match suggests that spacing of fractures, rather than fracture length, controls the formation of fine- and medium-sized fragments. On the other hand, the oversize coal is controlled by the fracture lengths, rather than the spacing. Data from subsequent test blasts support this idea. By increasing delay times from 8 msec to 17 msec, engineers were able to reduce fines to 7% - 10% but could achieve no further improvement, despite changes in shot pattern geometry, decking and hole size. The hypothetical fragment distributions show that, even without blast enhancement, about 5% of the natural coal blocks are fine-sized.

Jan 1, 1986

By B. Vasanthakumar, H. Ravishankar, S. Subramanian, J. Braun

The flotation of sphalerite and galena in the presence of intra- and extra-cellular components isolated from Bacillus circulans before and after adaptation to the chosen minerals has been studied. Based on various biochemical treatments, the key biomolecule primarily responsible for the flotation of sphalerite has been identified as DNA, which functions as a bio-collector. A noteworthy finding is the considerably higher flotation recovery of sphalerite obtained in the presence of single stranded DNA (79.8%) vis-à-vis double stranded DNA (50.2%). Selective flotation tests using the soluble fraction obtained after thermolysis treatment highlight that sphalerite-adapted cells yield a higher selectivity index (S. I.) of 31.2 compared to galena-adapted (S.I. 4.3) or unadapted (S.I. 12.2) cells. During bacterial adaptation to the above chosen minerals, the quantitative and qualitative changes in the surface potential, surface morphology and cell wall components have been monitored. The prior adaptation of the bacteria to sphalerite results in the secretion of a higher amount of proteins, eDNA and phosphate, whereas adaptation to galena leads to a higher amount of polysaccharide, uronic acid and acetylated sugar, vis-à-vis the unadapted bacteria. A significant observation is that the bacteria undergo a change in morphology from rod to sphere as a strategy to resist metal toxicity during adaptation to galena. Interestingly, this phenomenon is not observed during bacterial adaptation to sphalerite. This morphological transition has been shown to involve changes in crucial cell wall components as well as changes in the levels of expression of bacterial cytoskeleton element mreB gene involved in the maintenance of the rod shape. The sequence of B.circulans mreB gene has been identified using Polymerase Chain Reaction (PCR) followed by DNA sequencing.

Jan 1, 2014

By R. G. Sandberg

The Bureau of Mines investigated a hydrometallurgical procedure to recover gold, silver, and byproduct lead from a complex lead-zinc sulfide ore. A ferric chloride (FeCI3) preleach, at 40° C in 6 h, extracted about 50 pct of the zinc and formed insoluble silver and lead chlorides. Subsequent acid thiourea leaching, for 1 to 3 h at 40° C, extracted about 85 pct of the gold and silver. Lead chloride was solubilized from the remaining residue with NaCI brine solution at 55° C in 1 h. Lead metal was recovered from the brine solution in an electrolytic cell with an anion perms elective membrane at a current efficiency of 95 pct and current density of 100 A/m2. Spent FeCl3 leach solution was reoxidized in the anode compartment at a current efficiency between 75 and 95 pct.

Jan 1, 1986

By Y. Z. Zhang

Electrochemical oxidation of arsenopyrite in ferric sulphate media in the absence and presence of bacteria T jerrooxidans has been investigated for the purpose of understanding the dissolution chemistry of arsenopyrite and the bacteria attack mechanism. Mixed potential, polarization resistance measurements, polarization curve, cyclic voltammetry experiments were conducted in a ferric sulphate medium of 10g/L FeH, temp.25-35°C, pH 1.5-2.5 in the absence and presence of bacteria Tjerrooxidans, respectively. For the anodic process, a two-step reaction sequence is suggested by the electrochemical measurements. The initial step at a low potential is described by:

Jan 1, 1996

By Thomas F. Connors

The end of an era is here. The pony-express era of industrial relations in the coal industry is ending. Those professionals who know it must end to insure the industry&apos;s future are happy, relieved, but somewhat saddened. They are sad because it did have its elements of suspense, thrills, frustrations, and at times, plain fun. The days of the Company town are ended; the wildcats are declining; violence (while a big part of today&apos;s industrial relations) is beginning-to decline. The spirit of sound industrial relations is beginning to take hold. This has to happen for the industry to compete in today&apos;s economy and it is causing major changes for those who have grown accustomed to the pony-express type industrial relations.

Jan 1, 1981

By A. G. Matyas

Process improvements which have reduced the lead content of slag from the Brunswick lead blast furnace are described. Microscope examination and electron microprobe analyses, supported by the results of slag settling and viscosity tests, showed that typically about 70% of the lead in the lead blast-furnace slag was present in the form of entrained prills of metallic lead. The total lead content was found to increase with the magnetic susceptibility of the slag. Measurements in the furnace settler indicated the presence of a magnetite-rich layer between bullion and slag, arising from high magnetite levels in the slag. This layer was believed to inhibit the settling of entrained lead from the slag. Tracer tests showed that the average slag residence time in the settler was less than 2 minutes. Improvements in coke quality and sizing, which led to a reduction in the slag magnetite content, and the installation of a larger settler and a continuous Roy tapper, have all contributed to a significant decrease in the slag lead content.

Jan 1, 1977

By D. E. Macdonald

"Approximately 2.0 million tonnes of phosphate rock is imported from the U.S. to Alberta annually and is used in the production of agricultural fertilizers. No phosphate rock is presently mined in Canada, although sources of igneous phosphate have been mined in the past. Sedimentary phosphate rock is found in nine geological units within the Cordilleran Foothills and Front Ranges region of Alberta. The Devonian-Mississippian Exshaw Formation, Permo-Pennsylvanian Rocky Mountain Supergroup, and the Jurassic Fernie Formation have the best potential f or future development. Most of the phosphate rock in Alberta is of low grade and lies within restricted land use areas. The best over-all resource potential lies in the Fernie basin area of southeastern British Columbia, in the Fernie Formation. The possibility of deep phosphate rock deposits in the plains region should be examined.Phosphate rock is mined in Idaho, Utah, Montana and Florida, and is shipped to Canadian fertilizer plants for processing. Raw phosphate rock is acidified, usually with sulphuric acid, by a wet process. Varying amounts of ammonia are added, and the fertilizer is pelletized. Future research, in western Canada, is needed in the beneficiating of low-grade ores and in the area of in situ mining. At the present time, phosphate rock sells for around US$50Itonne, and Alberta fertilizer companies probably pay in the range of C$7D-$85Itonne for U.S. phosphate rock delivered to the plant. Environmental concerns about phosphate rock processing are centred around gypsum tailing ponds, and radioactive wastes.U.S. supplies of phosphate rock are secure until the year 2000; however, by this time, the richer deposits in the U.S. may be exhausted. Alternate deposits, such as the ones in Alberta and British Columbia, might then be considered. Phosphate byproducts, such as uranium, vanadium, fluorine and rare earth elements, may help in making the processing of low-grade Canadian phosphates economically feasible."

Jan 1, 1988

By Venugopal Tammishetti

One of the important technological challenges facing the iron ore industry in India is to find commercially viable technologies to process alumina rich iron ore fines and slimes. The key to solving this problem lies in developing selective reagents for the separation of iron oxide minerals from associated gangue minerals like quartz, gibbsite, kaolinite and montmorillonite. We present in this paper our successful results on the dispersion-flocculation of three different natural iron ore slime samples from mines in India using the reagent combination of PVP dispersant and corn starch flocculent. All the three samples contained hematite, goethite, kaolinite and gibbsite minerals as confirmed by X-Ray Diffraction (XRD) analysis. The effect of various process parameters such as pH, dosage of reagents, settling time and pulp density was studied. We were able to produce a concentrate assaying 66.2% Fe and <2% Al with a yield of 67.4% by weight (with a recovery of 77% of iron values) using PVP (1 ppm dosage) and corn starch (20 ppm) for one of the slime sample (Sample I) assaying 58.2% iron (Fe) and 3.8% aluminium (Al). The XRD analysis of settled and dispersed fractions revealed that alumina-containing minerals like clay and gibbsite mostly remain in the dispersed phase while iron-bearing minerals tend to concentrate in the flocculated fraction. Keywords: iron ore, slime, flocculation, dispersion, beneficiation, starch, polyvinylpyrrolidone

Sep 1, 2012