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By Ying Gu

Automated mineralogy methods and tools, such as the Mineral Liberation Analyzer (MLA) and the QEMSCAN, are now widely used for ore characterization, process design and optimization. Several case studies published recently demonstrate that large gains can be obtained through grinding and flotation optimization guided by automated mineralogy data. However, since automated mineralogy can only provide the information pointing to where the process gains can be made, it does not directly impact the production gain. Thus, a question is often asked: how to value the contribution of automated mineralogy to process improvement of a particular plant. This appears to be a difficult question. On close examination however, it is found that this is essentially a question of the value of information and it is reasonably well documented in various other industries. Hubbard (2010), in chapter 7 ?Measuring the Value of Information?, dealt with exactly this type of problem. The value of information is the reduced risk of an investment and opportunity loss. The methods Hubbard developed can be applied to estimate the value of automated mineralogy, as well as metallurgical test work, which also produces information that reduces the risk of investment. This paper first introduces Hubbard?s theory on the value of information and how to measure it. It then applies his methods to estimate the value of automated mineralogy, using Anglo Platinum?s fine grinding project as an example. In this case, the value of automated mineralogy includes the reduction in investment risk worth US$ 62.5million, in project rejection worth US$197 million and in opportunity loss worth US$320 million. In the end, a general model is developed to allow the simulation of the value of automated mineralogy in different mining operations constrained by different parameters.

Sep 1, 2012

By C. D. Harrison, J. W. Parkinson

This paper compares the ash slagging potential of Illinois No. 6 seam coal as raw coal, as commercially cleaned coal processed in a baum-jig plant, and as coal cleaned to a reduced ash specification at the Electric Power Research Institute's (EPRI) Coal Cleaning Test Facility. The comparison is based on ASTM ash fusibility data, a slagging index, gravity fractionation analyses of critical mineral matter compounds, and comparative pilot combustion tests. The first three grounds of comparison are laboratory data. The final comparison involves burning about 13.6 t (15 st) of each coal at Conoco's 4.40-kW/h (1.5 million-Btu per hr) test combustor. The evidence of these four comparisons indicates that lowering the ash specification of coal before combustion can significantly improve the ash slagging characteristics of Illinois No. 6 seam coal.

Jan 1, 1987

By T. M. Barczak

The term "resultant load vector" is defined as the representation of the forces applied to a longwall roof support element by strata activity through a single, quantifiable measure of support resistance. The relatively complex kinematics of the shield structure prohibits determination of support resistance simply from the summation of leg forces. In the research reported in this Bureau of Mines study, the mechanics of the shield structure were evaluated, and a technique was developed whereby the resultant shield loading could be determined by instrumenting supports with pressure transducers and strain gauges to measure leg, canopy capsule, and lemniscate link forces. This technique has been laboratory tested in the Bureau's Mine Roof Simulator. Functional relationships among variables were assessed, and confidence intervals were established for prediction of the resultant load vector parameters. Resultant load measurements were taken on five instrumented shields on an active longwall face in Colorado. Results were analyzed and found to be consistent with shield mechanics and anticipated roof behavior. Benefits to be derived from this research, future efforts, and long-range goals are discussed.

Jan 1, 1986

The following is a Maintainability Design Checklist for coal mining equipment. The purpose of the checklist is to provide a summary of design review points for the maintainability assessment of new or existing underground equipment. It specifically focuses on the identification of equipment design features, tasks, or procedures that impact equipment downtime, repair costs, labor hours and maintainer skill level requirements. Some checklist points are general in nature. The checklist was designed to be used across all categories of underground equipment. The intent is to draw attention to design features and maintenance procedures that will increase maintainability requirements. You are encouraged to adapt this checklist to site specific or machine specific requirements by: ? Inserting specific performance criteria for various categories of maintenance tasks. For example; all hydraulic lines on a shuttle car should be replaceable in 15 minutes or 25 minutes, etc. ? Adding or deleting checklist items for different categories of equipment. You would include environmental control equipment, for example, on face equipment and not on shuttle cars or man trips. ? Adding additional checklist items based on site or equipment specific maintenance histories or experience, company maintenance standards, or other factors.

Jan 1, 2000

By M. A. Mahtab

This report presents a new procedure for analyzing the orientations of rock fractures in an engineering site. This procedure, coded in a computer program, identifies clusters or groupings among the fracture orientations and calculates (1) the mean orientation of the fractures within each cluster and (2) the dispersion or scatter among these fracture orientations. Data points are represented by intersections of joint normals with the upper hemisphere, which is divided into 100 equal-area quadrilateral patches. The points in adjacent patches where the Poisson distribution test shows significant concentrations are assigned to a single cluster. For a cluster whose fracture orientations follow Arnold's hemispherical normal distribution, a confidence interval for the fracture orientation mean is calculated by applying Fisher's estimates. The technique is applied to the treatment of three examples.

Jan 1, 1972

By D. J. Dodds, W. L. Schroeder

INTRODUCTION The plate loading test is a simple, useful tool in rock mechanics. The test consists of applying load to a rock surface and noting the resulting deflections. Plate loading tests are typically used to obtain modulus of deformation and related material properties, but can also provide information on blast damage and in situ stresses (1). Properties obtained are of course dependent on test procedure and test environment . In order to eliminate extraneous results it is essential that the interrelationship between the physical test environment and various rock properties obtained be thoroughly understood. Test interpretation is based upon elasticity methods, and the general assumptions made during analysis which include the following: 1. The rock is homogeneous, isotropic, elastic and continuous. 2. The stress is induced in a semi-infinite solid. 3. The plate is perfectly rigid or completely flexible. 4. Rock properties are independent of in situ stresses. 5. The plate and rock surface are smooth and parallel.

Jan 1, 1974

By M. J. Virnig, B. D. Boley, G. A. Wolfe, J. M. W. Mackenzie

A flowsheet for the recovery of nickel and cobalt from laterite ores has been developed which involves acid leaching followed by hydroxide precipitation of the metal values and then releaching of the hydroxide cake with ammonia. Investigations of the chemistry of extraction of nickel by LIX® 84-I and a developmental reagent, XI-84-IT, along with the physical properties of the organic phases are described. Based on these investigations, suggestions with regard to design of a solvent extraction plant and operating conditions are made.

Jan 1, 1999

By Marcos Velho Ludolf

O projeto de Substituição dos Controladores Lógicos Programáveis da Sinterização tem como objetivo a substituição dos PLCs da Família S5 (Siemens) já obsoletos por novos PLCs de tecnologia atual. O projeto de substituição foi dividido em etapas, sendo que a primeira etapa ocorreu em 2012. Consequentemente, teremos novos e antigos PLCs na mesma plataforma de controle até o fim do projeto. Os maiores desafios foram fazer toda a migração e adaptação de hardware e software dos novos PLCs ao restante da planta adequando-se aos sistemas existentes tais como PIMS, comunicação com o Nível 2, comunicação com balanças já obsoletas, SCADA, outros PLCs e concluir toda a migração da primeira etapa do projeto em apenas 60 dias compreendendo desde a desmontagem do sistema anterior, montagem dos painéis novos, recomissionar todo o I/O substituído, conciliar todos os testes à frio e parciais com demais serviços de manutenção programados e realizar o comissionamento a quente de toda a sequência da planta nos últimos 04 dias de implantação, quando todas as linhas de operação foram liberadas para testes. Em 2014 foi iniciada uma segunda etapa, e um PLC de 600 pontos de I/O já foi substituído em apenas 80 h usando uma nova estratégia.

Jul 30, 2018

By F. J. Sotiuo

Variability of the quality of phosphate ore feed to the flotation plant has been of great concern for the industry. This variability sometimes resulted in 10% to 50% decrease of P2O5 recovery, a reduction of P2O5 concentrate grade and/or both due to depression of the system. For this investigation, samples of refractory ore and easy-to-float ore were obtained from the same mining area in Central Florida; and then submitted to bench flotation tests using fatty acids (FA). To understand the cause of difference in flotation behavior of these apparently similar ores (grade, mineralogy, and physical appearance), bulk and surface characterization studies were conducted. Based on the results of characterization studies and bench flotation tests on both refractory and easy-to-float samples, it was possible to determine FA surface coverage. It was found that the surface coverage for the refractory ore was between 6.9% and 9.8%, corresponding to 0% and 80% recovery, respectively; whereas the easy-to-float ore showed surface coverage of 17.2% (89% recovery) to 24.6% (98% recovery). Since size analysis, surface area and pore size measurements showed small differences, it was thought that the surface coverage differences could be caused by the ad-sorption of natural compounds onto the phosphate ore surfaces, depressing the system. SEM, EDS, and Leaching and Solvent Extraction (LSX) studies were performed on the phosphate ores to demonstrate the proposed mechanism. This paper summarized our characterization studies, flotation results, and findings. The study showed that the new identified compounds were responsible for the flotation behaviors of the Florida refractory phosphate ores.

Jan 1, 2014

By M. Nete

Qualitative analysis of three mineral samples from northern Mozambique was accomplished by X-ray diffraction (XRD), microscopy, scintillation, and magnetic determination. Chemical characterization was accomplished using ICP-OES and pressed powder XRF after dissolution using microwave-assisted digestion and flux fusion. The samples consisted of a mixture of minerals with tantalite as the major component, with minor amounts of microlite, quartz, mica, accessory garnet, and tourmaline. Incomplete dissolution was obtained using microwave digestion, with maximum recoveries of 90.25 ± 0.06% and 88.90 ± 0.04% for Nb2O5 and Ta2O5 respectively. Complete dissolution using Li2B4O7 resulted in recoveries of 98.5 ±0.2% Nb2O5 and 100.4 ± 0.3% Ta2O5.

Jan 1, 2012

By F van de Venter, J Monama, T Kale

For a multitude of years fine iron ore, -1 mm, across the world has been deposited as discard material in tailings facilities. This contributes to environmental problems, decreases plant yield and increases the cost of processing. Fine iron ore recovery processes are often deemed unfeasible or uneconomical, due to the complexity of the reclamation thereof in comparison with high-grade ores. However, recent advances in the recovery of fine iron ore using the Optima Classifier® has shown great promise. The Optima Classifier® is a new hindered settling separator that makes use of an upward current of water for density separation. It’s unique design combines three different separating zones for increased efficiency and recovery. In this study the potential of the Optima Classifier® to produce a +63 per cent Fe concentrate from -1 mm feed material was evaluated from a standalone and process perspective. It was found that a +63 per cent Fe product can be produced from a single stage, using the Optima Classifier®. Fine iron ore feed with an average head grade of 50 per cent Fe was upgraded to average concentrate grades of 63.2 per cent Fe. Displaying upgrade ratios of up to 1.26 in a single stage. A three-stage, rougher, cleaner and scavenger system did however show the best performance with recoveries of up to 70 per cent, yielding 56 per cent of a 63 per cent Fe product. Furthermore, the dewatering of the product and tailings was evaluated using the UltraG ultra-fine dewatering screen ensuring dry, stackable material. The tailings could be used in codisposal, while loads on thickeners and filter presses are significantly reduced. The high throughput of the Optima Classifier® ensures a small plant footprint, minimal civil and structural requirements and low operating utility consumption such as electricity, minimising the carbon footprint of a fines beneficiation plant. The Optima Classifier® in combination with the UltraG dewatering technology has provided a platform for economical recovery of fine iron ore, both current arising and historical. Providing opportunity to lower plant operating costs, reduce environmental liability of tailings facilities and increasing immediate process water recovery

Sep 18, 2023

By Royal Preston Jarvis

minerals of different specific gravities. It is simple in construction, easily operated, capable of treating large quantities in a short time, and highly efficient under various conditions. The question, whether the material to be jigged has first been sized, determines the two principal methods of jigging. Jigging preceded by close sizing, generally known as the Continental or German system, involves a more or less elaborate series of screens or trommels, with attendant cost for installation, operation, and repairs. Jigging without sizing, known as the English system, is, according to Munroe,1 a development of the hand-jigging formerly employed in Cornwall . and introduced by English miners to this country." In its simplest form, the method consists in jigging an ore-mixture previously crushed to some maximum size (although, in some cases, even this preliminary is omitted) oil a relatively coarse sieve, and then jigging again on a filler sieve, the material passing through the first sieve and bedding. While many modifications have been necessary to adapt it for use in mills of large capacity, where hand-work was necessarily, replaced by machines, the principle remains the same; the fact that the English system has been successfully employed, both in this country and abroad, is well known; and arguments have been made for its efficiency and applicability in a wider sphere than it has occupied hitherto. II. REVIEW OF PREVIOUS INVESTIGATIONS. The fact that treating a mixture of minerals under jigging conditions increased the amount of mineral saved, or, as Professor Richards aptly terms it, 1 ?the extra jig-catch," has long been known. To account for this fact a number of theories have been proposed. The work of Rittinger2 in this field has, for many years, been a classic in the literature of ore-dressing. For the purposes of my present paper, however, the work of two American investigators, Prof. H. S. Munroe and Prof. R. H. Richards, is chiefly concerned. Professor Munroe has given the results of an elaborate series of experiments,3 and his deductions, based largely oil theoretical grounds, of this work. After reviewing briefly the two systems of jigging, followed by a discussion of Rittinger's formulas and

Sep 1, 1908

By F. T. Cawood

Monitoring of the slopes of surface mines can prevent loss of life, loss of equipment, loss of production and possibly the loss of the mine. The effectiveness of such monitoring depends on the extent to which slopes give adequate advanced warning before failing, and on the ability of the monitoring system to detect such warning. In this paper an overview is given of slope monitoring requirements, from both the geotechnical and survey points of view, as well as the types of equipment and instrumentation that are available. Legal issues relevant to slope monitoring are also dealt with. It is concluded that conventional geotechnical and survey instrumentation will continue to have their place in slope monitoring. However, the slope monitoring scene has changed substantially in recent years. A real and volume-based monitoring equipment is now available and this has provided, and will continue to provide, the means to improve safety, and to enhance productivity in mining. Its benefit to slope management is substantial. In the near future it is expected that data obtained from such equipment will add a much needed new dimension to slope engineering, and be used to improve slope designs and to optimize slope angles.

Jan 1, 2006

By A. J. Naldrett

"Synorogenic mafic intrusions are a common feature of the northern Appalachians. Two sulphide-bearing intrusions in central Maine that have been investigated revealed a number of important features that illustrate sulphide-silicate interaction that relates to magmatic sulphides. The Moxie Pluton is a large mafic intrusion emplaced during the Devonian Acadian orogeny. The Katahdin gabbro is a small mafic body that is probably genetically and temporally related to the Moxie Pluton. Both intrusions were formed bv the intrusion of variably fractionated magmas into irregular chambers and feeder zones. The southern half of the Moxie Pluton contains three small sulphide accumulations, whereas the Katahdin gabbro hosts a large, predominantly massive, sulphide body. Textural evidence clearly documents the magmatic origin of these sulphides.The three Moxie Pluton sulphide bodies are located at or very close to the intersection between the intrusion and the sulphidebearing Hildreths Formation. Sulphur assimilation, suggested by this relationship, is supported by the isotopically light (634S = —8 to —25) and the selenium-poor (S/Se - 25 000-48 000) character of the Moxie sulphides and correlation with the Hildreths Formation sulphides (S34S = —25, S/Se = 25 000-48 000). Isotopically light sulphur (534S = —20 to —21) and high S/Se ratios indicate that the majority of sulphur in the Katahdin sulphide body was also derived from an unexposed country rock source. The volume of sulphide in this example illustrates the potential efficiency of the sulphur assimilation process—so efficient in this case that marked dilution of the economic chalcophile elements resulted from the addition of so much external sulphur.The three Moxie Pluton sulphide occurrences are all impoverished in Ni and Cu (1.87-0.44% Ni; 0.72—0.32% Cu in massive sulphide), but contain normal Co contents (0.22-0.13%) relative to typical magmatic sulphides. The Katahdin sulphides show a greater Ni and Cu impoverishment (0.2% Ni; 0.1% Cu in massive sulphide) with similar Co (0.17%). The Ni content of the Moxie Pluton olivines document the fractional removal of sulphides during crystallization of the host magma, which resulted in the progressive depletion of the chalcophile metals in this magma. Modelling that involved the fractional crystallization of the magma coupled with the fractional removal of a specified proportion of sulphide has successfully duplicated the Moxie Pluton sulphide compositions. A similar magmatic model has been postulated for the Katahdin sulphides with a greater degree of sulphide fractionation. This latter model is again supported by extremely low Ni/forsterite ratios in olivines, though the compositional range is insufficient to document fractionation. These magmatic models are not unique, but they provide a relatively constrained and consistent explanation for the sulphide compositions.The Ni contents of olivines are the most important indicator of sulphide-silicate interaction in mafic magmas and have exploration potential. They provide a means of assessing the timing and extent of sulphur saturation in an intrusion and are a guide to the Ni content of a sulphide accumulation in equilibrium with olivine. In the Maine examples they provide strong evidence of the magmatic origin of the sulphide accumulations and justification for magmatic modelling."

Jan 1, 1984

By J. Alton Burks

When acoustical materials are used in a mining environment for noise control purposes, they are subject to environmental deterioration from hydraulic fluid, moisture, and dust. These and other factors can cause physical degradation of the material, which can lessen its ability to absorb sound. In this study, the Bureau of Mines measured the sound absorption properties of 16 different acoustical materials after 4 sample treatments: (1) being kept clean and dry (as received from the supplier), as a control or reference standard, (2) immersion in water and draining, (3) immersion in lOO-pct-petroleum-type hydraulic fluid and draining, and (4) exposure in a coal mine. The last three treatments were used to approximate the type of physical degradation that can be experienced in actual use. The impedance tube or standing wave method was used to measure the normal absorption coefficient. It was found that the absorption coefficient of most materials was adversely affected by the retention of either oil or water, with oil having the greater effect. The only class of materials that was affected by neither oil nor water was neoprene foam.

Jan 1, 1990

By Vladimir Adamek

This paper describes the Bureau of Mines development of a novel subsidence prediction methodology suitable to the mining and geologic conditions in the Northern Appalachian Coal Region. It describes the computation of vertical and horizontal movements, inclination, curvature, and horizontal strains. The substance of this method is the separation of the effects of lithology by introducing a correlation between hypothetically homogeneous overburden and existing lithologic conditions, while providing for different mining conditions, i.e., underground geometry and overburden thickness. The effects of lithology have been expressed in the form of a variable subsidence coefficient within the subsidence trough. The subsidence co-efficient is considered a constant for other predictive methods. Field data from 16 test sites at 11 Bureau longwall panel studies were used in the analysis. For each panel, the characteristics of the variability of the subsidence coefficient were defined. Regression analysis of the subsidence coefficients from all test sites on their locations relative to the edges of their respective panels yielded a third-degree polynomial equation. The results from additional longwall panel studies, not included in the regression analysis, were used to prove the validity of this method. To facilitate the use of this precalculatlon methodology, a computer program was written in BASIC for use on a personal computer.

Jan 1, 1987

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core. For this reason neither analysis of core recovered (amounting to 37.5 pct of approximately 10,000 ft of hole drilled) nor sludge analysis considered separately would give correct assays. The following equation has been adopted: actual weight theoretical actual weight core — theoretical weight sludge Drilling is done un.der contract. Collection of core and sludge samples together with double-checked assaying is done by the company. Microscopic, specto-graphic, and X-ray examination of rock specimen saved is carried on at the University of Utah. A fund, set up by Kennecott Copper Corp. at the University, has encouraged cooperation in solution of mining research problems. The drill used is a Longyear hydraulic powered by a Waukesha 26-hp gasoline motor. A pump, Bean Royal type, is powered by a 7-hp Briggs-Stratton gasoline motor. Weight of unit is 4000 lb and is placed under tripod of 35-ft wood poles within a 25-ft sq area. The following results have been attained: Avg depth of hole drilled, ft 1215 Avg drilling depth per 8-hr shift, ft 10.11 Core Size, In. Avg depth of NC casing, ft 21 2 11/16 Avg depth of NX casing and size core, ft 178 2Ye Avg depth of BX casing and size core. ft 427 1% Avg depth of AX casing and size core, ft 585 1 5/32 Through the use of metal sludge collection boxes, 12 ft long, 18 in. wide, and varying from 12 to 24 in. in depth, sludge recovery has been maintained above 95 pct. Samples, both core and sludge, are assembled for each sample interval and averaged in 50-ft bench increments. Determinations for mineral values are made at both the mine and mill. Assay results varying in excess of 0.05 of copper assay values are rerun to determine correct assay. The organization chart for the Kennecott Utah copper mine is shown in Fig. 1. The Utah mine requires approximately four men per shovel shift to drill solid rock, properly slope banks, and trim loose material from the top of bench sections in making them safe for men and equipment working below. During recent years, while the mine has been worked day and night, powder crews have worked day shift only. Series of toe holes drilled back of advancing electric shovels comprise approximately 70 pct of all mine drilling. These holes, spaced about 21 ft apart, have a depth of 24 ft. Of the footage drilled, 25 pct is for vertical holes either 24 or 28 ft deep, spaced about 28 ft centers along the top rim of banks. This combination of toe holes, using an Ingersol-Rand DA 35 drifter mounted on a tripod and the vertical holes using an Ingersol-Rand Type X71 wagon drill is standard practice above the pit elevation. These benches are 68 ft high, with a few exceptions, while benches below the pit elevation are 50 ft high and require very little drilling other than toe holes. Five benches, varying in height from 80 to 100 ft, require a combination of toe holes plus vertical holes and bank holes or electric churn drill holes. Bank holes using the DA 35 drifter and electric churn drill using Bucyrus-Erie Model 29T, account for 5 pct of feet drilled. Until the past year, drilling of fill material on upper levels was accomplished by churn drill method. Wagon drills are now used for fills because they can be set up and drill a hole an hour. It is necessary to run an air drill 7 hr for each 8-hr shovel shift. Average drilling rates during the past year were as follows: DA 35 drifter using 2 and 3-man crews — 4.36 ft per man hr X 71 wagon drill using 2-man crews — 6.45 ft per man hr DA 35 drifter bank holes, 3-man crews — 2.53 ft per man hr Bucyrus-Erie churn drill, 2-man crews —1.9 ft per man hr Total mine avg — 4.57 ft per man hr Mine practice assigns powdermen to job locations where they perform all drilling and blasting work within the area. Almost as much labor is consumed in springing and loading holes as in drilling them. Average powder consumption is approximately as follows: Per Electric Shovel Shift, Lb To spring and chamber holes — 50 Blasting — 500 Trim banks — 50 One fifth of all powder used is Hercules hercomite bag powder while four fifths is Hercules gelamite 65 pct strength in 1 Vi in. x 8 in. sticks. Frozen material, overhanging from top banks, creates a winter hazard to shovel operators and equipment. Experiments are being conducted with a high-speed Hercules powder in an effort to eliminate this condition. Banks are trimmed by powdermen who lodge 8 or 10 bundles of powder, containing 15 or 20 sticks each, across the upper surfaces of each working face. To each bundle of powder is attached a primer with 6-ft fuse which is lighted as the powdermen, who have lowered themselves down the bank on ropes, ascend to safety. Three angle dozers, D8 caterpillars, make grades and move supplies for the average 31 drilling machines working daily. Compressed air is furnished by three Ingersol-Rand compressors, each with rated capacity of 3750 cu ft of air per min at 100 lb reservoir pressure.

Jan 1, 1952

By Thomas A. Palmer

Free cyanide has been successfully determined in ground and surface water by isolation using a microdiffusion technique followed by a pyridine-barbituric acid photometric finish. The determinable free cyanide includes hydrocyanic acid, ionic cyanide, and the cyanide of weak complexes such as cadmium, lead, manganese, and zinc cyanide. Complexes of copper, iron, nickel, and silver are sufficiently strong so the dissociation into free cyanide does not occur at the microdiffusion pH 7. Cyanate, sulfide, and thiocyanate do not interfere. Samples containing as little as 0.005 mg/L are successfully analyzed. The relative standard deviation is <10%. Samples stored at pH >I2 in the dark are stable at least three weeks.

Jan 1, 1981

By Y. F. Liang

Fe-6.5wt.%Si alloy has excellent soft magnetic properties compared with conventional silicon steel. It is very brittle at room temperature due to appearance of ordered phases, like B2 and DO3. In this paper, we investigate the effect of ordering on cold workability of this alloy. Methods of rapid quenching and hot rolling were utilized. It is found that rapid quenching and proper heat treatment of the hot rolled sheet could decrease the degree of order due to the very high solidification rate and high cooling rate. The rapid quenched sheet and hot rolled sheet after proper heat treatment could be cold rolled without crack. Microstructure, phase ordering, micro hardness, and cold rolling properties were investigated by Optical Microscope, X-ray Diffraction, Micro Hardness Indenter and Four-high Mill.

Jan 1, 2009

By H. B. Brunot, Ilsley. L. C., E. J. Gleim

"INTRODUCTION One of the various phases of the work by the U.S. Bureau of Mines, in its endeavors to minimize the hazards that attend coal mining, is the investigation of electrical motor-driven machinery. The object of this investigation is to foster the development of explosion-proof machines for use in gassy mines. This investigation is conducted under the provisions of Schedule 2C, which prescribes the inspection and tests made upon motors, controllers, and other electrical parts to determine their ability to withstand internal explosion without discharge of any flames. The method of conducting these inspections and tests is described in Bureau of Mines Bulletin 305, Inspection and Testing of Mine-Type Electrical Equipment for Permissibility, published in 1929. Formal approval of a machine for use in gassy mines is not granted by the Bureau of Mines until the machine is found to comply with the requirements, including inspection and test. When it is approved it is entitled to classification as a ""permissible"" machine.The high regard in which machines approved by the U.S. Bureau of Mines are held by engineers of other countries is attested by the growing number of shipments of permissible equipment; to other nations, including Australia, Russia, Norway, Japan, and Manchuria."

Aug 1, 1936