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By A. J. S. Benade, N. B. Strydom, W. H. Van Der Walt

The influence of two weeks of microclimate cooling on the state of heat acclimatization by N. B. STRYDOM, M.Sc., Ph.D. (Illinois) (Fellow), A. J. S. BENADE, M.Sc. (Potch.) (Visitor), and W. H. VAN DER WALT, B.Sc. (Potch.) (Visitor) Five volunteer Black subjects were acclimatized according to the climatic-room acclimatization procedure, and their states of heat acclimatization were determined by the usual heat-stress test. The subjects thereafter wore pre-frozen jackets in a somewhat higher environmental temperature (33,9/35,6 °C). stepping at the same rate as before for four hours a day for fourteen days. The heat-stress test was repeated and followed by a similar test at the higher environmental temperature. It was found that fourteen days of microclimate cooling in a very hot environment had no adverse effect on the degree of heat acclimatization. The sweat rates during microclimate cooling remained high, even though no sweating is possible beneath the pre-frozen ice-jacket. which indicates that there is an increase in sweat rate from the areas of exposed skin.

Jan 7, 1975

By J. F. Perreault, J. Paraszczak

"The profitability of an underground mine depends to a great extent on the production cost.Random equipment failures and their consequences have significant influence on the total operating cost of a mining system. Having this in mind, a reliability study has been initiated for loadhaul- dump machines in an underground mine. The main purpose of this study is to identify the critical machine subsystems, determine their failure pattern and help work out an effective maintenance strategy. In the first part of this study, the failure data of diesel powered LHDs, gathered during an I8-month period have been analyzed. For the purpose of the analysis, the machines have been treated as systems comprising seven subsystems connected in series. The most critical subsystems, i.e. electrical and hydraulic, were further investigated, and analyzed by graphical and analytical methods. Different failure distribution models were applied and discussed. Basic Defmitions and ConceptsIn order to avoid confusion, several terms used in this paper are defined below, as compiled by O'Connor (1991). The acronymsBS and MIL-STD stand respectively for: British Standard and US Military Standard.Failure: the termination of the ability of an item to perform a required function (after BS 4778).Time Between Failures (TBF): the length of time between two consecutive failures.Observed Mean Time Between Failures (MTBF): For a stated period in the life of an item, the mean value of the length of time between consecutive failures computed as a ratio of the cumulative observed time to the number of failures under stated conditions (after BS 4778).Maintenance, Corrective: The actions performed, as a result of failure, to restore an item to a specified condition (after MIL-STD-721 B).Maintenance, Preventive: The actions performed in attempt to retain an item in a specified condition by providing systematic inspection, detection and prevention of incipient failure (after MIL-STD-721B)."

Jan 1, 1994

By R. G. Schaal

The Magma Mine at Superior, Arizona had an unimpressive beginning as a. worked out silver mine that was purchased in 1910 for $130,000 and then incorporated into the Magma Copper Company which has been continuously mining the deposit since 1915. Copper is the principal metal produced. However, zinc, gold, silver and iron have also contributed to the total metal product. Prior to 1973, the production rate of 1500 tons per day was maintained by hoisting 4 and 5 ton pay-loads at speeds up to 1100 FPM through four small hoisting shafts. The hoists used included 8' and 7' diameter double drum, double clutched doc. and a.c. motor driven geared units built as early as 1917. One old timer that did yoeman's service for many years was an 8' Bollen Pulley with a.c. geared drive that utilized an endless rope system with mechanical grip and release mechanism built into the pulley's throat. Primary power for all of these old production hoists was 25 Hertz. Exploration that was completed in 1967 developed a reserve of 10 Million tons of 5 per cent copper which was sufficient to justify new capital expenditures for enlargement of productive facilities to 3300 tons per day. To reach the new reserves a 22' I.D. concrete lined vertical shaft was sunk to an initial depth of 48431. A 12 ½ diameter adit was bored from the mill side of the mountain to intercept the shaft at a point about 1100' below the collar. This 9200' long adit provides the right-of-way for the mine to mill railway and egress for heavy supplies from rail head as well as a conduit for potable water, electric power, compressed air and mine drainage. Mine operating levels begin 3582' below collar and thereafter have been cut at 100' vertical intervals. Skip loading stations have been installed 4282' and 46821 below collar with an in-shaft skip dump station located 800' below collar. This provides for 4500 tons of storage above the mine to mill train loading station located 1100' below collar. Ore and waste are moved horizontally from stope chutes and drift headings by rail haulage to transfer raises which channel all material to the two mentioned skip loading stations. The 22' diameter shaft crossection is arranged with the two skip compartments on the south side separated from the manway and utility compartments by the cage compartment. The cage counterweight operates in a portion of the southwest quadrant of the shaft. The hoisting plant consists of two conventional Nordberg double drum, single clutch, geared, d.c. motor driven units that are identical except for the drives and automatic controls. The drums are 15' diameter x 105" face with LeBus grooving for 2-1/8" diameter wire rope wound in three layers. Mechanical design emp1o::ed is the same as has been used for the past twenty- five years. One drum is attached to the main shaft by 2 large bolted integral shaft flenge. The loose drum is engaged by a 4-arm positive tooth clutch meshing with radially cut teeth of the clutch ring which is bolted to the drum. The drums are one-piece steel plate weldments having a finished weight of 60 tons each and are mounted upon a 36" diameter shaft that also weighs 60 tons. With the addition of brake rings, clutch, bull gear and rope, the total dead weight adds up to nearly 250 tons that is supported by just three double row spherical roller pillow block bearings. There is no bearing between the drums. All pinion shafts as veil as all head sheave journals are supported in the same type of antifriction bearings.

Jan 1, 1975

By S. Zhang, Z. Hua, D. Feng, X. Tong, X. Xie

The separation and beneficiation of low grade collophanite has long been a technical challenge in phosphorous production. It requires the joint efforts of process design, reagents optimisation and operational control. In this study, the statistical response surface method was applied for determining the optimum reagent dosing in the flocculate flotation of a low grade (18.59%) collophanite ore. The dosage of three reagents, including a pH regulator (Na2CO3), depressant (Na2SiO3) and flocculant (modified starch), were input parameters of the modelling design. Two quadratic mathematical models were derived for prediction of grade and recovery based on the statistical design. Statistical analysis of the experimental results validates the reliability of the models. Through further calculation, the optimum dosage of the three reagents were found to be 3,020g/t Na2CO3, 1,990g/t Na2SiO3 and 880g/t modified starch, with which a maximum recovery of 75.81% and P2O5 grade of 30.08% was obtained from flocculate flotation test. Keywords: Process optimisation, modelling, froth flotation, collophanite ore, response surface methodology

Jan 1, 2020

By B. K. Choi, G. E. Kim, S. Noh, T. K. Kim

Oxide dispersion strengthened (ODS) steel has been developed as the most prospective core structural material for next-generation nuclear systems because of its excellent high strength and irradiation resistance. In ODS steel, finely dispersed nano-oxide particles with a high number density in the homogeneous grains can be very attractive to achieve superior mechanical properties at high temperatures, and these favorable microstructures can be significantly changed using process parameters during the mechanical alloying procedures. In this study, the effects of mechanical alloying times on the microstructure of 10Cr-1Mo ODS martensitic steel were investigated. Pre-alloyed powders with Y2O3 powder were mechanically alloyed for various milling times using a high energy horizontal ball-mill apparatus (model: symoloyer CM20). Powder properties such as the mean particle size, size distribution, and crystallite size were evaluated. Uniaxial hot pressing was employed to consolidate the milled powders. The results of a microstructural observation after the MA and UHP processes indicated that the homogeneity of the martensitic structures was considerably increased with the longer milling time. It is thus concluded that the mechanical properties of the ODS martensitic steel are greatly dependent on the mechanical alloying times.

Jan 1, 2015

Environmental leaching methods such as the toxicity characteristic leaching procedure (TCLP) test or the ASLP test are used to assess wastes including mining and mineral processing by-products. These tests were initially designed as regulatory tests simulating conditions in a co-disposal municipal landfill waste facility and do not allow leaching performance to be predicted in site specific scenarios. Further, the assumptions underlying these tests tend to define worst-case scenario leaching conditions for by-products from the mineral industry disposed of outside a co-disposal scenario.   Recently some characterisation leaching test methods have been issued as ISO/TS standards (ISO/TS 21268-3:2007 and ISO/TS 21268-4:2007) that aim to define the leaching performance of soil type wastes and to predict the leaching performance of these wastes over long time frames. In this preliminary work these standards are trialled for by-products from the mineral industry and the quality of information that can be obtained from them compared to TCLP/ASLP tests.   Where clear comparisons were possible, the TCLP/ASLP tests were typically found to extract a similar, or often greater, amount of component than (i) tests conducted at the same pH but without the acetate buffer or (ii) for test conditions aimed at simulating long-term leaching in an eluate defined by the waste type. While further work is required, the preliminary results indicated that the tests may provide reliable long-term estimates of leaching performance of wastes in soil types of variable pH and provide a suitable basis for long-term planning for optimal waste handling, treatment or even re-use.

Jun 1, 2010

By H. Ewaschuk

"INTRODUCTIONRio Algom's involvement with the Enviro-Clear thickener began in May 1975, when tests on a laboratory waif were carried out by our Research and Development group. Commencing in June, further tests were carried out using a 3 foot diameter pilot unit. By December, a 30 foot full-scale thickener had been installed. In the near future, 3 more Enviro-Clear thickeners will be commissioned. With in the space of a year, almost $2.5 million had been committed to liquid/ solid separation facilities using the Enviro-Clear thickener. This paper briefly traces the extent and development of this commitment.BRIEF HISTORY OF QUIRKEQuirke Mine started operating in September 1956 at a rated capacity of 3,000 TOD. Because of market conditions, Quirke was shut down in December 1960. The facilities were re-opened at an increased capacity of 4,000 TPD in July 1968, by which time the demand for uranium had returned. Also, a second shaft was sunk, and by 1970, all of the ore treated in the concentrator was coming from the new Quirke II mine. The Quirke I stopes were allowed to fill with natural seepage. As well, acidic barren solution from the concentrator was pumped underground.With increasing demand for uranium, the capacity of Quirke will be further expanded to 7,000 TPD by 1978. Studies of how this expansion could best he carried out led to the consideration of a fairly new piece of equipment, the Enviro-Clear thickener, for certain operations in the concentrator. These operations are as follows:"

Jan 1, 1977

NIOSH recommends that exposures to respirable coal mine dust be limited to 1 mg/m3 as a TWA concentration for up to 10 hr/day during a 40-hr workweek, measured according to current MSHA methods (see Section 5.1 and Appendix J). NIOSH recommends that sampling be conducted with a device that operates in accordance with the NIOSH accuracy criteria [Busch 1977; Busch and Taylor 198 11 and the international definition of respirable dust [ACGIH 1994; CEN 1993; IS01993; Soderholm 1991a,b; 1989].* The REL represents the upper limit of exposure for each worker during each work shift. For single, full-shift samples used to determine noncompliance, NIOSH recommends that MSHA make no upward adjustment of the REL to account for measurement uncertainties [NIOSH 1994c] (see also Section 5.6.2). NIOSH further recommends that all reasonable efforts be made to reduce exposures to respirable coal mine dust below the REL through the use of engineering controls and work practices.

Jan 1, 1997

By M. A. Stogran-Baker, S. Cole

"Fluorspar mining occurred in the St. Lawrence area from the 1930s until the mid 1980s. In 2013, CFI discovered a new ore body, the AGS vein, which proved amenable to more economic open pit mining methods. A new mill was designed to produce 200,000 tonnes per year of high quality, acid-grade fluorspar concentrate. The simple mill flowsheet includes crushing, dense medium separation (DMS), grinding, sulphide flotation, pulp thickening, fluorite flotation, thickening, and concentrate filtration. The type of flotation cell used is being installed for the first time in North America. Site clearing commenced in the spring of 2016, followed by mill building erection completed by the fall of the same year. Civil and mechanical installation as well as site-wide construction was completed by the end of September 2017. Commissioning commenced, and at the time of paper submission mill production is on track to ramp up to full capacity by the beginning of 2018. This paper will focus on the use of fluorine NMR analysis applied to fluorite and the development of on-stream analysis, flotation cell selection and the decision to install the new cells, and the water control program designed to maintain process control with the least possible environmental impact.HISTORY St. Lawrence is located on the south-eastern coast of the Burin Peninsula in Newfoundland and Labrador. The first occurrence of fluorspar in the St. Lawrence area was recorded in 1843 by British geologist J.B. Jukes (DNR 2010). Fluorspar was likely extracted in small quantities prior, however commercial fluorspar mining did not commence until 1933 at the Black Duck Mine. The Black Duck Mine operation processed ore by washing and screening – acid-grade was hand-picked from the oversize, the undersize was sold as a lower grade product. The Black Duck Mine was closed in 1941 (Wilson 2010). The American Fluorspar Limited (New Fluor) commenced operations on the Director vein in 1936, and was bought by Alcan in 1942 (Cooper 2012). Alcan also developed the Iron Springs as well as the Tarefare deposits. The Black Duck Mine mill had closed, and a new mill was built. This new mill crushed the ore prior to dense medium separation, the sinks of which were sent to Quebec for use in Alcan’s aluminum smelter (CFI 2016). Alcan shut down it’s mining operations in 1977 due to the dropping price of fluorspar as well as labour disputes. In 1986, the British-based Minworth Limited began producing ore from underground at Blue Beach North and Blue Beach South, as well as from open pits at Tarefare and Director veins (Cooper 2012). A new mill was built by Minworth in order to process the ore using flotation to produce an acid-grade concentrate. The low price of fluorspar caused Minworth to shut down operations in 1990. Whilst exploration has been ongoing, no mining had occurred in the area since 1990."

Jan 1, 2018

By Verne C. Fryklund, S. Warren Hobbs

The Coeur d'Alene district in the panhandle of Idaho is one of the major lead-zinc-silver producing areas in the world. The value of recorded production to date has exceeded $2 billion. Country rock comprises six formations of the fine-grained, highly siliceous, and slightly metamorphosed Precambrian Belt Series. Small stocks of monzonitic rocks and several varieties of dikes intrude the sediments. The district is at the intersection of a north trending major anticlinal uplift and the west northwest-trending Lewis and Clark line of which the Osburn fault is the dominant feature. The district structure is complicated and comprises folds, some of which are overturned, and a complex of faults of diverse ages and directions of movement. Late Cretaceous stocks but are also probably Cretaceous in age. The productive veins are localized within a series of 12 subparallel, straight mineral belts that trend approximately N65°W and that appear to be controlled by deep fractures in the basement rocks. The mineral belts are in two groups, separated by 15 miles of post-mineralization movement on the Osburn fault. Veins and ore shoots are variously disposed within each mineral belt. Many crop out at the surface, but others apex several thousand feet beneath it. Local controls that influence the position of ore shoots within the mineral belts include the brittle character of host rock attitude of bedding relative to the mineralized fracture, and the amount and direction of movement on faults. Depth appears to have had little effect on the occurrence or type of ore. Ore occurs in a series of steeply dipping replacement veins of relatively simple mineralogy. Six periods of mineralization, ranging from Precambrian to Tertiary in age, are recognized. The productive galena sphalerite-bearing and tetrahedrite-bearing veins-of the main period of mineralization-are younger than the

Jan 1, 1968

By G. Marchand, R. H. Hay, E. C. Pearson

"AbstractA polishing .technique is described together with two anodic etching methods for aluminum and its alloys. Some of the optical and structural characteristics of the film formed by •these etching methods are discussed and it is shown to what degree these properties of the film are related to the crystallographic orientation of the metallic substrate. Applications to the study of recrystallization and twinning phenomena and to the study of plastic deformation, banding, and texture estimation are discussed. Micrographs are shown .to illustrate the manner in which the technique is particularly well suited to the study of such phenomena.IntroductionThe physical metallurgist is constantly faced with the problem of polishing and etching of metallic sections. Mechanical polishing is tedious, time consuming, and not applicable whenever a surface entirely free from plastic deformation is required. Etchants are generally required to reveal the grain structure in order to determine its size, shape, and regularity, or to bring up constituents. In the case of aluminum, the .most common etchant to reveal grains is a mixture of three acids: HC1, HN03, and HF. This etchant is fairly satisfactory on certain alloys and for certain metallurgical conditions but in general fails to give a good definition of the grain on castings, on highly alloyed products, and whenever the grain is smaller than 100 microns in diameter. Besides size and shape, a grain, or crystal, has an orientation and may be endowed with several types of imperfections."

Jan 1, 1952

By N. G. Macfarlane, G. C. Howell, C. R. Mtshali

This paper explores the philosophical risks of role misalignment in tailings storage facilities governance. Central to this discussion are the roles of the Engineer of Record, Responsible Tailings Facility Engineer, and Accountable Executive as required by the international standards like the Global Industry Standard on Tailings Management. The paper examines how these roles interconnect. It critiques current governance practices and calls for a shift toward genuine engagement.

Apr 8, 2026

Identifying Opportunities to Reduce the Consumption of Energy Across Mineral Processing Plants

Sep 13, 2010

By Sabine Verryn, Johan de Villiers

"Through the application of the Rietveld method, X-ray diffraction analysis has developed into a fully quantitative method for phase analysis. The Rietveld method itself has progressed from an unstable method that required constant operator intervention to a stable and robust method suited even for routine analysis. The method however, requires a good initial calibration and periodic checks on its precision and accuracy.Several examples from minerals processing, pyrometallurgy, and materials science show the strengths and applicability of the method. The limitations and sources of error are also discussed.1. IntroductionDiffraction of X-rays by matter has been used for more than 80 years, initially to elucidate the structure of crystalline matter, and to investigate the bonding between atoms. The first development of powder diffraction cameras and the later development of X-ray diffractometers saw the development of methods to identify phases and to try to quantify them in mixtures. A comprehensive database and search-match techniques have been developed over the ensuing years. Effective and accurate quantification had to wait for the so-called Rietveld' analysis where the total diffraction pattern is used and where the crystal structure data for the constituent phases are needed as input. Overlapping of diffraction peaks of different phases is handled without difficulty and reliable results can be obtained in spite of this occurrence.The development of methods for structure analysis from powder data has also progressed rapidly and more and more complex structures are being solved from first principles, i.e. without knowledge of unit cell or space group data. Because of the ease of use of the Rietveld method, it is finding increased application in metallurgy, especially in the quantification of phases of importance in minerals processing, in pyrometallurgical processes, and in materials characterization. This study will examine the principles of the Rietveld method, its application in metallurgy, and finally also the common pitfalls that need to be avoided.2. Rietve1d AnalysisAs the Rietveld method compares the experimental diffraction pattern with one calculated from known crystal structure data, the following data are needed:•,An XRD Pattern with as Iowa background as possible. This should have good resolution and high diffraction peak intensities. With modem X-ray detectors, this can be achieved in minutes rather than hours.•,Instrumental parameters such as goniometer radii and slit sizes are needed. These are the so-called ""Fundamental parameters"". They are used to calculate the shapes of the resultant diffraction peak profiles that are caused by the instrument. The parameters are used to correctly separate the instrumental effects on the diffraction pattern from the sample effects."

Jan 1, 2014

By Erwin L. Zodrow

"This paper is based on a statistical analysis of the Smallwood iron ore mine (6). In this mine, the prediction of magnetite is especially difficult because of strong trends. Recognizing the statistical limitations of the sample data, the paper approaches the basic statistical analysis of the magnetite by endeavouring to obtain a first approximation of its sample density function. The knowledge of the distribution of a random variable is essential for several reasons, but particularly: (1) to obtain reliable estimators of the population parameters -for example, mean and variance; and (2) to use these estimators in hypothesis-testing concerning future experiments -sample points and sample intervals."

Jan 1, 1968

By W. D. Musgrove, B. W. Gandrud, A. C. Richardson

"INTRODUCTION The Cahaba coal field is the second largest coal—producing area in Alabama. This field underlies parts of Shelby, Jefferson, St. Clair, and Bibb Counties and has an area of about 270 square miles. The Cahaba formations contain a number of coal beds, about 8 of which are mined at present. Although the area contains many folds and faults, samples from various mines and beds indicate that, except for the variation in ash content, the coal from this field is very uniform in character and quality. It is a high-volatile bituminous coal used chiefly for steam-raising purposes and domestic fuel. The coal as mined usually contains too high a percentage of impurities to be marketed without some preparation. The majority of the mines have washing plants in which jigs are the favorite type of equipment, although at some mines tables are used for cleaning the slack sizes. The washing characteristics of the coal from the Cahaba field are very uniform and differ from those of the near-by Warrior field, in that the coarser sizes contain a much lower percentage of bony material than the finer sizes. In general, the bone increases with decrease in size. It breaks into thin flakes, making difficult its elimination in the cleaning process, and gives rise to what is known among the washery operators of the Cahaba field as the ""rash problem.""A previous publication 5 referred to rash from the standpoint of jig operation and suggested some means of handling it with jigs. This report presents the results of table tests on two samples of fine coal from the Thompson and Woodstock beds, respectively, of the Cahaba field to determine the influence of rash on table efficiency.This investigation has been made in connection with the study of coal-washing problems now being conducted by the U.S. Bureau of Mines and the University of Alabama, with the cooperation of the different coal operators of Alabama."

May 1, 1934

By S. Lolon

Spontaneous combustion in underground coal mines has become a serious problem particularly in the caved area (gob). Recent statistic has shown that approximately 17% of total 87 underground coal mine fires in the United States are attributed to spontaneous combustion (De Rosa, 2004). Spontaneous combustion results from a self-heating process in exothermic conditions. The accumulated heat, if not removed, is conducive to the rapid increase of temperature and may result in mine fires or explosions. It is well accepted that the interaction between oxygen and coal substances is the main cause for spontaneous combustion, while other factors such as pyrite, moisture, and bacteria play a secondary role to the self-heating of coal. Therefore, only coal oxidation is considered in this study. Coal oxidation occurs as coal comes into contact with air. This process involves complex phenomena in terms of heat transfer, chemical surface absorption, and energy balance related to inherent properties of coal (Wang et al., 2003). For simulation purpose, overall reactions can be simplified as suggested by Mitchell (1990): C + O2 CO2 + heat [65°- 94°C] (1) CO2 + C 2CO + heat [100°- 150°C] (2) Stoichiometric reactions (1) and (2) show that 2.66 grams of oxygen are required to oxidize 1.0 gram of carbon. Once the temperature exceeds 100°C, the reaction has little chance to stop. The crucial step in reducing spontaneous combustion risk is locating the ignition point of spontaneous combustion (hot spot). This information is useful in the effort of developing a preventive method effectively. In this study, this was conducted based on the best gathered information: ventilation surveys conducted in an existing longwall mine located in the western US; the laboratory experiments performed on a physical gob model; and the CFD models to investigate the flow behavior in the gob, the oxidation of coal, and heat transfer phenomena. The hot spot location is determined as a function of oxygen concentration and gob temperature. The critical values are the following: 5% (by volume) for oxygen and 100°C for gob temperature. CHARACTERISTICS OF GOB MATERIAL AND MODELS Permeability Concept Permeability is one of the key parameters in any study dealing with porous medium. This is determined by particle size and the ratio of void volume to the total volume (porosity) of porous medium. This relationship is known as the Carman-Kozeny equation (Scheidegger, 1957) and is given mathematically as [ ] where k* = theoretical specific permeability, m2 dm = the mean particle size, m n = porosity However, permeability tests with various particle sizes were carried out at the University of Utah?s ventilation model (Figure 1) to verify this relationship. These tests indicated a necessity to modify this theoretical relationship. A factor of 0.898 was obtained to make up the different between theoretical permeability and the experimental one. The modified version of (3) is given by: [ ] This equation is used to determine permeability of the simulated mine gob throughout the study. [ ] Mine Gob Material The gob is represented by zones filled with material of given size distribution. For physical measurements and CFD simulations, the material used were crushed rock and coal. Due to compaction, three different characteristics of gob material were used in the model: unconsolidated, semiconsolidated, and consolidated. In the real condition, the largest coal-rock particles are more likely to be located in the area behind the shields. This material is freshly broken and unconsolidated. The size of these broken particles is based on a study carried out by analyzing images taken from the area behind the shields in three coal mines in the United States (Pappas and Mark, 1993). The results have shown that the mean size of materials behind the shields is about 1.22 m. The permeability associated with this size was then obtained from (4) and permeability tests. This was 4.68 x10-7 m2. For semiconsolidated and consolidated zones, due to the lack of experimental data, these were determined through CFD simulations. These simulations were designed by assigning a permeability of 3.15 x 10-8 m2 for consolidated zone and 7.98 x 10-9 m2 to the other zone. The airflow pattern in the gob assigned with given permeabilities was expected to follow the experimental distribution (Brunner, 1985). The mean particle sizes for the semi-consolidated and consolidated zones were 0.02 and 0.006 m, respectively; smaller than those of the unconsolidated zone. Figure 2 shows the 3D view of permeability changes in the gob.

Jan 1, 2009

By R. Altindag

The rock brittleness is one of the most important rock properties that affect the drill-ability of rocks. It is supposed that the increase in rock brittleness causes the increase in penetration rate. The drill-ability of rocks depends on many factors. These factors can be separated as two parts depending on a machine, bit type, rotational speed, thrust, blow frequency, flushing, and rock properties, brittleness, compressive strength, structure, hardness, etc. In this paper, the relationship between brittleness and drill-ability index, which is used in the prediction analysis of the penetration rate of rotary blast hole drills, were examined using regression analysis. It is also seen that significant correlations exist between brittleness and the drill-ability index.

Jan 1, 2002

By C. I. Walker

Field tests at a number of large mineral processing operations have been undertaken to better understand the wear life impact of different aspects of pump design, materials of construction and maintenance procedures. Data are presented for wear rates, and typical result variability is explored. To improve the availability of pumping equipment in a milling circuit it is not only important to improve the wear life, but also reduce the variability in the wear life of components so that the incidence of premature failure is lessened. Localised gouging wear due to flow separation and turbulent eddy formation is often the point at which part-holing occurs. A number of improvements to the design of pump impeller and throatbush (suction side liners) to reduce the localised gouging wear are outlined. Also, results are presented on the use of composite materials (mix of metal, rubber and ceramic) for throatbush materials to protect specific areas against wear.

Jan 1, 2012

By H. E. Strauss

The time-to-fracture test has been applied to pure platinum and to two alloys of platinum under the special conditions of small cross-sectional area of the specimens and of a test temperature above the annealing temperature. While the results are similar in character to those secured with large specimens of steel at temperatures below the annealing temperatures, two differences are noted: (I) an upward inflection in the curves for log stress vs. log time for the alloys, and (2) a sharp break in the curve for log stress vs. log time for pure platinum when single crystals filled the wire. The experiments showed that: (I) platinum and 10 per cent rhodium-platinum maintained their strengths at IIOO°C. better than did to per cent iridium-platinum, and (2) the rhodium-platinum after a time exceeded the iridium-platinum in strength at IIOO°C. Platinum alloys in their service at high temperatures encounter conditions somewhat similar to those that exist in the usual form of the "time-to-fracture"l test with steel. This paper is intended primarily to present results obtained by a "time-to-fracture" test under conditions approximating actual service for platinum (at the temperature IIOO°C. under oxidizing conditions) with three materials, pure platinum (99.99 per cent), a platinum alloy with 10 per cent iridium, and a platinum alloy with 10 per cent rhodium. The experimental procedure differed from that previously used in this type of measurement in that the samples were of small diameter (0.010 in.), were fully annealed, and were tested at a temperature definitely above their annealing points (600°C., IOOO°C. and 1050°C., respectively, for 60 per cent cold reduction and 15 min. heating time2,3). The experimental arrangement was similar to that previously described,' using as heating unit a small, horizontal, platinum-wound furnace, ½ by 12 in., maintained at constant temperature by an automatic controller. The samples were 0.010-in. wires, which had been annealed at 1200°C. by drawing them slowly through the furnace. The time-to-fracture was measured electrically. Elongation was not determined because of the difficulties inherent in the experimental arrangement. Life—that is, the time-to-fracture—was considered the important property in these first tests. No attempt was made to measure the reduction in area, but the types of fracture were noted qualitatively; in the text they are referred to as (I) the normal, plastic, or necked fracture, (2) the knife-edge fracture, and (3) the brittle or intercrystalline fratture with small reduction of area. Results for different constant applied stresses are plotted in Fig. I. In all cases at least two trials were made for each stress, although they cannot always be indicated separately on the curves. The breaking Stresses drop very rapidly to a small frat-286

Jan 1, 1943