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By R. J. Wasilewski

THERMAL expansion of rhenium data have been reported by Agte et al.,&apos; and Medoff and cadoff,&apos; respectively, while the linear expansion coefficient was determined by Sims et Al.3 Denoting expansion coefficients parallel and at right angle to c axis by all and Al respectively, their weighted mean should be equal to dilatometric linear coefficient of a random polycrystalline specimen. The values reported for these coefficients are tabulated below: Since these values are in poor agreement, and because of anomalous expansion behavior observed in titanium 4 and chromium, a rede termination of the axial expansion was carried out. A Unicam 19-cm-diam high-temperature camera was used, with unfiltered CuK radiation. Camera calibration was carried out using at Pt wire standarde at temperatures up to 1100oC, the variation between indicated and true temperature being within the experimental error (estimated at <3oC) above 750o C. Rhenium Johnson Matthey spectrographic standard) was vacuum annealed at 1000 oC prior to X-ray diffraction experiments. Lattice parameters were obtained by least-squares solution using the reflec- tions (1232), (1015), (2024), and (3030) for all the patterns, and the additional reflection (2133) for higher temperature (400°C and above) patterns. The relative expansion data obtained are shown in Fig. 1, lengths of the vertical bars indicating the spread in the parameter values obtained by using various combinations of three spacings from the four or five measured. The data calculated from the equation of Ref. 2 are included for comparison purposes. Several specimens were used to obtain the above data, neither of them showing a significant parameter change after completion of the high temperature runs. Some parameter data and the corresponding expansion coefficients between 25 oC and the measurement temperature are tabulated below. The comparison of the present data with those previously reported shows reasonable agreement with the results of Medoff and cadoff2 within the temperature range covered by them (up to 1000 oC). The a values are appreciably higher throughout most of the investigated temperature range. The basal expansion is—in agreement with their findings—appreciably higher than that in the direction of hexagonal axis. The expansion coefficient, al, however, decreases rather more slowly with temperature than previously reported. The expansion along the c axis up to nearly 1000°C appears in excellent agreement. However, at higher temperatures there is a marked deviation towards higher expansion, reproducible between different specimens. Furthermore, there seems to be some irregularity in the variation of all with temperature, the apparent local minimum and maximum values at 510o and 1024o C, Table II, being well outside the experimental error limits. It must be stressed that the apparent discontinuity in the c parameter expansion is still insufficient to account for the expansion value reported by Agte et al.&apos; since a total expansion of close to 2 pct would

Jan 1, 1962

By T. A. Klinefelter

Lightweight concrete aggregates are materials weighing less than the usual aggregates of sand, gravel, and crushed rock. Concretes made with sand and gravel or crushed rock weigh 145 to 150 lb per cu ft, compared with 50 to 90 lb per cu ft for concretes made with lightweight aggregate. The ultra-lightweight concretes weigh much less than 50 lb per cu ft. Lightweight aggregates may be natural or manufactured materials, or byproducts from other commercial operations. Natural materials include pumice and pumicite, tuffs, breccia, scoria, diatomite, and vegetable products, such as peat, straw, and sawdust. Manufactured materials include expanded perlite, vermiculite, slag, clay, shale, and slate. Byproducts include cinders, aircooled slags, and coke breeze. Concretes made with lightweight aggregate were used first in construction of commercial buildings, later in homes. Present usage is general in multistory and other building construction, such as bridge decks, barges, jet-plane runways, piers and other structures. The industry has grown remarkably in the last decade. General Considerations The particular use to which the concrete may be put determines the importance of the various properties of the aggregates employed. Any given type of aggregate, however, should be reasonably uniform in composition and properties. It should be suitably graded and maintain proper weights, insulation values, and adequate strength to prevent breakdowns in mixing. It should bond properly with the cement, be inert to chemical reactions with cement and steel, and resist weathering, moisture, insects, and fungi. Some properties of any given type of lightweight aggregate from different suppliers may vary considerably because of differences in the source of raw material and the method of treatment. This statement applies particularly to expanded clays, shales and slates. All lightweight concretes have higher absorption of moisture and lower thermal conductivity, lower compressive strength, as well as lower unit weight as compared with standard concretes made with sand, gravel or crushed rock. Excepting cinders, the cost of lightweight aggregates is usually higher. Grading and unit weight requirements are given in ASTM standards: C 330-53T, Lightweight Aggregates for Structural Concrete; C 331-53T, Lightweight Aggregates for Concrete Masonry Units; C 332-54T, Lightweight Aggregates for Insulating Concrete. The production of masonry units shows the growth of the use and types of lightweight aggregates. In 1947 the total number of lightweight block was 460,000,000 units in 8 by 8 by 16-in. equivalents. By 1954 this had increased to 950,000,000 units. Of the 1954 production, 27 pct was of manufactured types, 43 pct of cinders, and 30 pct of remaining types. Raw Materials Used as Lightweight Aggregates It was stated above that a variety of products both natural and manufactured are used as lightweight aggregates. It is easy to understand that a bulky material of low unit value cannot be transported far and still remain a low cost product. Hence the product used in a particular market depends upon the availability of the possible raw materials. Thus it is

Jan 1, 1960

By A. Keil

TO measure the absolute conductivity of metallic specimens it is generally necessary to have exact geometrical shapes and good electrical contact with the test-piece. To overcome limitations of the standard methods, an electronic conductivity-meter* Trade name, Sigmatest; model Dr. F. Forster; Reutlingen, Ger-many. has been developed, which works on the principle of inducing high frequency eddy-currents in the specimens and measuring the reaction of these curients on a measuring coil.&apos;,&apos; To attach the instrument to the test specimen, a flat area of about 10 mm diam is all that is necessary. The reading is taken off a scale calibrated in pet of International Annealed Copper Standard. The induced eddy-currents are propagated in a circular motion parallel to the surface. Thickness of the layer included in a measurement depends on the frequency used. With a frequency of 60 kc per sec, the penetration of the currents was 0.8 mm at the upper limit of the scale, and 2 mm at the lower limit. The method is limited to nonferrous metals and other applications have been described elsewhere." The object of the present investigation was to detect the formation and growth of surface layers of differing conductivity, and to determine the depth to which these changes could be observed. A well known phenomenon which illustrates the problem is the internal oxidation of certain silver alloys containing base metal solutes, such as copper or cadmium, when they are annealed in an oxidizing atmosphere.4, 5 Until now, the only methods available for following these phenomenon have been metallo-graphic examination and hardness tests.0, 7 Changes in mechanical properties of copper alloys by internal oxidation were recently discussed in a more general way by Martin and Smith." When measuring these conductivity changes it must be remembered that the only absolute values of conductivity are those of 1) the nontreated homogeneous specimen, and 2) the final values obtained with a surface layer of depth comparable to the penetration of the eddy-currents. All intermediate values refer to the mixed conductivities of the transformed surface layer and the untransformed alloy, and are thus readings of relative value only. Experimental Results Specimens were annealed for 1/2 hr at 800°C in hydrogen, and the internal oxidation was carried out by thermal treatment in air at 800°C. Fig. 1 shows the results of the first series of experiments. With internal oxidation of a solid solution of 8 pet Cd in silver, the conductivity rises continuously until the surface layer formed has become thick enough to be the deciding factor (about 400µ). Precipitates of CdO are mainly in the form of points, Fig. 2, and, as a result, the electric contact between the silver crystals is not distorted by them. Brine11 tests were run with a 1.25 mm ball and a 7.812 kg load. The specimen annealed in hydrogen and oxidized to a depth of 400µ had a hardness of 40 Bhn,

Jan 1, 1959

By L. G. Schulz, P. Spiegler

The electrical resistivity and the temperature coefficient of resistivity were measured with a potentiometric method using pure mercury as a reference material. Measurements were made nea,v roorn temperature. A detailed study was made of Hg,TI, near its melting temperature. THE work to be described was originally undertaken to provide the necessary data for interpreting the results of some optical experiments,&apos; but because of the growing importance of liquid metals,273 it appeared that the results of the resistivity measurements should be published separately. For convenience the alloys selected were those which are liquid at room temperature. Special consideration was given to the Hg-T1 alloy closely corresponding to Hg,Tl, because recent X-ray diffraction results of Smallman and Frost seemed to show that the solid structure may persist to some extent into the 1iquid. EXPERIMENTAL PROCEDURE A potentiometric method was employed in which the resistivity of mercury was used as a reference. The essential features of the samples are shown in Fig. 1. A capillary tube CT with a bore of approximately 1.5 mm was fitted into two end cups, T, which were made of Teflon. A pair of current electrodes CE were immersed into the excess liquid metal which partly filled the end cups. The potential electrodes PE were inserted into CT through short pieces N of a no. 18 hypodermic needle held in place with Saureisen cement SC. As shown, the lower ends of the PE7s were reduced in diameter to 0.25 mm to minimize the distortion of the liquid metal cross section. Several methods were used to insert an alloy into the sample, but the usual one with mercury alloys was to fill the sample container with a weighed quantity of pure mercury* and then add weighed the sample tube in a vacuum. Because of rapid corrosion, samples containing gallium were of necessity prepared in a vacuum. To obtain the specific resistivity of an alloy the resistance of the sample containing the alloy was measured and compared with the resistance when the container was filled with mercury. Resistances were measured in the temperature range of 20" to 60°C to obtain the temperature coefficient of resistance. The electrical measurements were made by connecting the sample in series with the standard one-ohm resistance, and the potential drop measured alternately over the sample and the standard resistance. To correct for stray thermoelectric effects provision was made for reversing the direction of the current, and the true potential drop was taken to be the average of that obtained for the two directions of current flow. Various materials were used for the electrodes, the most uniform results being obtained with nickel. EXPERIMENTAL RESULTS The numerical results are presented graphically in Figs. 2 and 3 and in tabular form in Tables I and 11. Following are remarks on results of individual alloys. 1) Hg-In—This was the first alloy studied, and

Jan 1, 1960

By Paul J. Sirkegian

A detail of the guide rollers is shown on Fig. I; and Fig. 2 shows a photograph of a guide roller that has been in service for several months on the haul-back cable. The former practice of supporting and guiding the haul-back cable on slusher hoists at Kimberly was accomplished by using the conventional type of snatch-block pulley, but it was found that an excessive amount of cable whipping and unreasonable wear on both rope and pulley resulted from this type of cable support. The guide, which is wholly made in the Corporation&apos;s shops, consists of two vertical and one horizontal roll assembled in a frame that is lagged to the cap timber in a slusher drift. In some slusher drifts more than one guide-roller assembly is used, depending upon the length of the drift. The installations usually are placed at 30-ft. intervals. We have found by actual experience that these rollers have increased the life of the haul-back cable approximately 200 per cent, as compared with the usage of the conventional sheave-block pulley. Cable Care The underground rigging crew and the slusher operators are trained in rope splicing. No cable clamps or knots are permitted on slusher cables and all cable "eyes" are spliced. The standard practice has been to use preformed cables of ?-in. diameter for haulage and ½-in. diameter for the haul back. The preformed cables have proved to be important factors for safety in use and speed in making splices. No external lubrication is used on the cables. It has been found that dry ropes have about 400 per cent longer life than those that were lubricated under these operating conditions. An accurate record is kept of all cable life and the tonnage moved by the slushers in each working place, and if the rope wear is excessive in any particular working place, an investigation is made to determine the reason. Wire-rope Sling for Snatch-block Anchor The rope sling is shown by Figs. 3 and 4. These slings support the snatch-block pulley for the haul-back rope. The standard practice at Kimberly is to use a roller-bearing snatch-block pulley 10 in. in diameter. The snatch-block housings are equipped with a clevis attachment in preference to swivel hooks, as it has been found that the more rigid clevis keeps the cable in better alignment. The snatch blocks are hung from the cable sling, which is wrapped around the posts of the drift sets as shown on Fig. 3. This method of anchoring the pulley provides better anchorage and flexibility than those of the rigid or spring-mounted types. When the slusher-drift headings are being advanced, the snatch blocks are hung from a "gin" pole ahead of the last set of timbers&apos; An employee is assigned on each production shift to take care of the lubrication of all slusher hoists that are being used and he is provided with tools for minor repairs.

Jan 1, 1946

By John W. Brauns

18.1-RAILROAD JOHN W. BRAUNS Conventional railroad haulage can be employed at both open-pit or strip and underground mines-in the latter case, largely for transfer of ore from distant openings to processing plants, or processed material to loading points or docks, using standard hopper cars for bottom unloading or gondolas for rotary dumping. The major application, however, is in surface mining for moving ore and/or waste out of the pit to a crushing plant or dump. Consequently, the following will deal largely with this phase of railroad operation, though the conclusions apply equally, for the most part, to other rail-type surface haulage. In surface mining, trains consisting of a locomotive and gondola or side-dump cars suitable for loading by power shovels are used. Gondolas are restricted to hauling ore as a rotary dumper is required. Side-dump cars are more versatile, carrying both ore and waste, but because of their additional dumping mechanism they are likely to be more expensive to maintain.&apos; Also they weigh up to twice a flat-bottom car of the same capacity and. therefore, fewer" cars can be hauled per train. Standard-gage track is most commonly used. Track systems incorporate low-quality temporary track where moves are frequent and permanent high-quality track on main lines. Rail haulage can be used alone, or in combination with other surface haulage methods, thus permitting a selection best adapted to the peculiar requirements of each mining project. 18.1.1---GENERAL APPLICABILITY, CONTROLLING PHYSICAL CONDITIONS For handling rough rock, the shovel-train system is unexcelled.&apos; The ideal applica¬tion is in a large, but not excessively deep, open pit mine from which the coarsely blasted ore and waste must be transported over 3 mi. If. in addition, a high production rate is to he maintained, railroad haulage probably will afford a lower cost per ton transported. Grade must he relatively gentle-usually 3% uphill and 4% downhill. The greater the differences in elevation involved, the greater are the track grade space requirements; for example, a 3 c ramp out of a pit only 300 ft deep is almost 2 mi long. Main-line track locations should he selected carefully so that the number of moves during the life of the mining operation can he kept to a minimum. This applies both to the ore haulageway and that for waste. Such moves are not only expensive, but are difficult to execute without adversely affecting production rates. The large mine affords greater opportunities for judicious location of semipermanent trackage. Figs. 18-1 and 18-2 illustrate typical track layouts. The effect of terrain in considering railroad haulage is confined chiefly to the initial cost of the installation. Grade construction in mountainous territory is ap¬preciably more expensive. However, if that expense can be borne, the subsequent maintenance cost of trackage will not he much higher than in areas of low relief.

Jan 1, 1973

By Matthew Anderson, David Day, Paul Smith

"The Midwest is presently the focus for Combined Sewer Overflow (CSO) abatement in the United States (U.S.). Most of the EPA mandated utility and municipal long term CSO abatement programs in large metro areas in the Midwest such as Cleveland, Indianapolis, Columbus, Chicago, St Louis, Akron, Fort Wayne, Louisville, Kansas City, Omaha and Pittsburg use or will use deep tunnels for storage and conveyance of CSOs. This paper will compare design methods specified and implemented for tunnel construction in Midwestern geology such as soft ground and mixed face EPBM tunnels, hard rock TBM tunnels, drill & blast tunnels, road header tunnels along with tunnel lining systems. It will also review the performance of successful CSO systems that include design life cycle, vortex and drop shaft configurations, pump stations, storage and conveyance strategies and long term operation and maintenance criteria. HISTORY and BACKGROUND Prior to development of mandated wastewater treatment plants, collecting both wastewater and storm water runoff in a single-pipe system was an efficient and economical solution. The early combined sewer systems were typically designed and sized to carry several times the average dry weather wastewater flows and not designed to handle a mixed volume of waste water and storm water during a wet weather event. Beginning in the early 1970s, federal legislation (Clean Water Act of 1972, Clean Water Act of 1977 and the Water Quality Act of 1982) changed how municipalities and industries in the U.S. were required to treat and discharge wastewater into the nation’s surface waters. They were mandated to treat wastewater at federal regulatory standards for discharge into surface waters that is safe for humans, wildlife and aquatic life. The wastewater treatment facilities that evolved were initially designed and sized to treat only the peak volume of raw wastewater flowing during dry weather. Combined sewer system control sections or relief structures were installed upstream of wastewater treatment plants (WWTP) and were designed to protect WWTPs from potential damage that would be caused if peak volumes of raw wastewater mixed with storm water runoff reached and overwhelmed the WWTPs."

Jan 1, 2016

By R. H. Clarke

"INTRODUCTION Characterization of waste rock is frequently an oversight in mine planning which can contribute to significant delay in permitting and ultimately development. Thorough waste rock characterization is now a prerequisite to developing operating and closure plans acceptable to regulators and the public. BACKGROUND A long history of mine development has placed virtually all focus on the economic extraction and beneficiation of ore materials, while waste rock characterization and management received substantially less attention and consideration. Inadequate waste rock characterization has contributed to planning and management practices that are having unintended and unforeseen impacts to surface and groundwater, and in some cases lead to significant long term closure liability. In response, the regulatory community has developed guidelines for performing waste rock characterization during the pre- mine permitting process that in some cases seem onerous (BLM, 2013). Additionally, where a project proponent has not met these guidelines or results have been deemed “inadequate”, the Nevada Division of Environmental Protection and Bureau of Land Management (regulatory agencies) will only issue operating permits with rigid management practices to address the uncertainty, including requirements for lining waste rock facilities. WASTE ROCK CHARACTERIZATION FOR NEVADA PROJECTS Newmont Mining Corporation (Newmont) operations in Nevada approaches waste rock characterization like ore characterization, as it requires a similar investment of time, funding and attention to detail in order to pass public and regulatory scrutiny. Newmont follows a tiered approach to characterization that creates three over-lapping data sets representing an “escalation” in analysis at each tier, an approach that has thus far gained regulatory approval. “Tier 1” waste rock characterization for most Newmont projects begins with exploration drilling which follows internal standards and procedures. 20-foot composite samples are prepared from all drilling and analyzed for multiple trace elements using inductively coupled plasma mass spectrometry (ICP-MS) methods and for acid-base characterization using method ASTM-E1915. If the exploration program is successful and further resource or reserve development is recommended, the Tier 1 data are used to prepare a Waste Rock Characterization Plan (Plan) that will be presented to regulatory agencies as a part of baseline project characterization."

Jan 1, 2015

By Fred D. Wright, Stefan H. Boshkov

12.1-BASIC AND PARAMETRIC CRITERIA IN THE SELECTION, DESIGN AND DEVELOPMENT OF UNDERGROUND MINING SYSTEMS STEFAN H. BOSHKOV AND FRED D. WRIGHT The criteria governing the selection of an underground mining system relate to several factors the importance of which varies with geographic location, the state of development of the industrial technology, the economic climate in the Nation and, perhaps, in the world-all these as functions of time. Thus, the best choice of a mining method depends upon sets of definable and undefinable values for numerous factors that may be considered fixed at the precise point in time. Subsequent changes in economic and/or technological developments may necessitate modification or complete revision of the existing mining system. A classification of criteria for selecting a mining system must be based on the broadest definitions. Specific cases may present simple solutions or pose very intricate questions, the answers to which are frequently based on judgment rather than explicit logically reasoned and derived results. Specific criteria may be of negligible consequence in some instances and of overriding importance in other cases. It would be foolhardy to attempt to point out, in the limited space available, all possible variations and interplay of factors, as well as to define the setting in which this interplay can take place. The reader is, therefore, referred to the case histories described in the subsequent parts of this section where actual practice will point out the variability of the problems encountered in the selection of a mining method. A popular thesis suggests that the proper approach to selecting the method of mining should be rooted in the inverse solution, whereby the limitations of the various methods in use serve as criteria to eliminate most of these methods from consideration in the specific case. This narrows the ultimate selection to one or perhaps two most appropriate methods. At this stage, a critical comparison analysis will lead to the ultimate choice. Although such a negative approach un¬questionably is valid in discarding obviously inapplicable mining systems from consideration, it falls short in defining the best system, which in practice quite often turns out to be a variation of a textbook standard, or a combination of two or more such standards. Because the deployment of capital and manpower over a long-term commitment closely follows the selection of the mining system, the decision-maker can well afford to spend a great deal of time and effort in evolving the optimum solution for the peculiarities of his specific case. Within the basic engineering tenets of economy and safety to which the design of a mining system must adhere lie numerous factors of constraint which may be grouped, somewhat arbitrarily but for ease of discussion, into several categories. The criteria that enter into the choice of a mining method derive from the following groups of factors: 1. The shape, size and spatial position of the ore deposit.

Jan 1, 1973

By M. S. Badollet

"DURING the past five years, production of asbestos fibers has increased in all countries that list asbestos as one of their mineral resources. Despite the many small mines that have been opened in Canada, United States, Mexico, South America, Italy, South Africa, Australia, Yugoslavia, and Rhodesia, the supply still does not meet the demand. It is estimated that total 1952 world production of asbestos fibers of all varieties approximated 1,600,000 tons, with Canada supplying 966,382 tons (I), or 60 per cent of all fiber mined.Asbestos fiber brokers continually are offering odd-lot samples of chrysotile, amosite, crocidolite, tremolite, and anthophyllite fibers to manufacturers, •quoting fantastic prices for tonnages •varying from just a few to 50 or more in a month. In many cases, the samples are not properly graded and contain quantities of rock impurities and finely ground dust.When vendors submit samples of asbestos fibers, the manufacturer should consider• them in terms of price, quality, quantity, and as possible substitutes for fibers currently used in his products.Grading of these odd-lot fibers in most cases has been poorly done, with resultant non-uniform fiber lengths. In some instances, a number of miscellaneous lots of fibers have been .blended for sale. This blending produces a fiber that is difficult for the manufacturer to accept as a substitute for the well-standardized Canadian fibers."

Jan 1, 1953

By Jr. Hazen

Two statistical techniques, the mean-square-successive-difference test and correlation, are used to investigate the problem of assigning an area of influence to an assay. Data from five hydrothermal deposits, Climax molybdenum, Cebolla Creek titaniferous iron, San Manuel copper, Piedra Hueca copper, and Butte copper, are studied. Two groupings of the assays are used for the mean-square-successive-difference tests. One group involves an increasing sample volume by reaveraging assays to represent a longer sample intervals, the other by regrouping assays to represent longer sample intervals without a change in sample volume.

Jan 1, 1967

A series of physical tests have been conducted to determine the actual performance of bunton to shaft side wall connections typically used in new shaft construction in South Africa. The test set-up is described, together with the applied loading, which incorporates a cyclical load component and a quasi-static load component, in an attempt to simulate actual conditions in the shaft. The test results are given and it is shown that generally the performance is adequate, provided double-slotted holes are not used. Generally the fatigue performance is found to be good, with the only exception being cases where the bottom of the bunton is lower than the seating level for the bunton to side wall connection. Arising from the tests, typical design values of connection stiffness, strength and fatigue life are proposed and a design procedure including these three key aspects of connections is described.

Jan 1, 2005

By K White, R Hardie, J Merritt, B Kirsch, R Lucas

Constructed stream diversions are a common element of mine sites across Australia. A review of constructed stream diversion performance in the Bowen Basin funded by the Australian Coal Association Research Program (ACARP) in 1999 found over 80 per cent of constructed stream diversions failed to meet regulator and operator expectations. Since that time, improvements have been made in the assessment, design, monitoring and management of constructed stream diversions. Fundamental to these improvements has been the development of a better understanding of the regional stream systems and efforts to replicate the natural stream processes in the function and operation of constructed stream diversions.This paper discusses the history of constructed stream diversions through the Bowen Basin in central Queensland recorded in the outcomes of a series of ACARP funded research projects. Further, the paper reports on the outcomes of a current ACARP funded project that has reviewed the success or otherwise of constructed stream diversions in the region. The assessment has compared the performance of diversions that have been designed to replicate and provide for ongoing stream geomorphic processes with those diversions that were designed and built prior to the development of these design improvements. Based on these investigations we have identified design and construction successes and issues requiring further research and management attention.CITATION:White, K, Lucas, R, Hardie, R, Merritt, J and Kirsch, B, 2013. Challenges and opportunities for constructed stream diversions in the Bowen Basin, Central Queensland, in Proceedings Water in Mining 2013 , pp 73-82 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 26, 2013

By A. C. Johnson

This paper is one of a series being prepared by the Bureau of Mines on exploration, development, mining methods and costs at various mines in the United States. It describes such methods at the Lincoln tungsten mine of the Wah Chang Mining Corp., Lincoln County, Nev. The purpose of these reports is to promote development and conservation of our mineral resources by disseminating technologic information on mining methods, practices and costs. The Lincoln mine in Lincoln County, Nev., in the southeastern part of the [tate], was acquired by the Wah Chang Mining Corp., in April 1951. Mining methods and costs are of special interest at this property because of the low costs and good production record achieved through a well planned program of exploration, development, and exploitation. Operations were discontinued in July 1957 after the Federal Government suspended the tungsten purchase program.

Jan 1, 1959

By S. S. Wang

Dialkyldithiophosphinates are a new class of flotation collectors for base metal sulfide ores. Structurally, they are quite different from the conventional collectors such as mercaptans, xanthates, thionocarbamates and dithiophosphates. Collectors Structures [Mercaptans R-SHS Xanthates R-OCS-M+ Thionocarbamates R-OCNH-R Dithiophosphates 2 R-CCNH-R&apos; Dithiophosphinates (R) 2-PS-M+ ] In the above structural formula, R and R&apos; stand for a hydrocarbon radical, and M+ is an alkali metal or ammonium cation.

Jan 1, 1982

By N. C. Joughin

A simplified economic model is derived for the financial return from a new gold mining project. The model is used to analyze the effects of changes in the mining rate, capital costs, working costs, value of the reef and the stoping width. The most effective avenues for improving the return from a new project are indicated. It is shown that those technological innovations which would improve the return significantly, and therefore make it possible to open more mines, are the use of smaller mines, means for improving the mining rate to capital costs ratio for smaller mines, planning and control of mines by computer simulation, an improved mine valuation procedure and means for mining at a decreased stope tramming width.

Jan 1, 2014

By G. A. Munoz

Gold leaching of an auriferous pyrite ore from the Portovelo deposit in Ecuador indicates that noncyanide leaching systems, such as thiocyanate and thiourea, are not as effective as the traditional cyanide system. For example, whereas approximately 95% of the gold can be leached from the ore with cyanide, only about 50% gold dissolution can be achieved with a thiocyanate leaching system. Electrochemical measurements indicate that gold is not cathodically protected by pyrite in these systems. Based on the results of preliminary diagnostic experiments, it appears that gold dissolution in thiocyanate and thiourea systems is limited by passivation due to the formation of either elemental sulfur or sulfur compounds on the gold surface.

Jan 1, 1999

In this paper we present the results of preliminary investigations into the applicability of new laboratory scale tests for the prediction of operating gap and specific throughput in high pressure grinding rolls (HPGR). The work is a part of the CAMIRO HPGR research program being conducted at the University of British Columbia, which aims to develop a suite of small-scale tests and procedures for the selection and sizing of HPGRs. A focus of the presented study is to test sample types in terms of their terra-mechanics based characteristics and relate them to specific throughput and gap results from pilot HPGR testing on the same samples.

Jan 1, 2012

By N. Sinclair

"In open pit dozer bulk push operations, the material left resting against the highwall after blasting poses an issue. This is because dozers cannot physically get in behind the waste to begin the bulk pushing process. Instead, an excavator can be used to side-cast this waste for the dozers to access. An alternative technique is to have the dozers ‘side-cut’ the material, which involves the dozers running parallel to the wall, digging out the material, and eventually rotating their blades to begin bulk pushing. Industry practices suggest the latter technique is less productive and economical than using an excavator. However, such excavator sidecasting ‘economic viability’ assumes that smaller/ancillary type excavators are used (i.e. less than 23 m3 bucket capacity). So, what happens if some mine sites are restricted to using relatively larger excavators (> 23 m3 bucket) that are preferred for higher priority digging tasks? When can they avoid using their excavator? This paper explores multiple scenarios that have been designed and developed to simulate data results, which identify conditions in which the dozer side-cutting technique becomes more economically viable than excavator side-casting in terms of decreased unit cost per loose cubic metre. IntroductionOptimizing the cost of waste removal operations is important for the profitability of opencast coal mines. In the past decade, dozer bulk pushing has emerged and taken over from traditional waste removal operations such as dragline and truck-and-shovel methods at various sites around the world (Beatty, 1994). In a typical bulk push operation, once the bench has been blasted, waste is left resting against the highwall, which can either be cleared using a loading unit or by having the dozers ‘side-cut’ the highwall.Dozers perform most effectively when slot dozing side-by-side, perpendicular to the highwall. However, when dozers are required to remove material off the highwall, they must push parallel to the highwall, rather than perpendicular, hence ‘side-cutting’ the material to ensure a clean face. While dozer sidecutting is possible, it has been found that the majority of mine sites use a loading unit in this situation to avoid dozer inefficiency. Other perceived issues with dozer side-cutting include higher maintenance requirements and safety risks to dozer operators working in close proximity to the highwall."

Jan 1, 2019

By A. Adeli, A. V. Metcalfe, E. Addo, E. Sepulveda

"The accurate modelling of geometallurgical data can significantly improve decision-making and help optimize mining operations. This case study compares models for predicting copper recovery from three indirect test measurements that are typically available, to avoid the cost of direct measurement of recovery. Geometallurgical data from 930 drill core samples, with an average length of 19 m, from an orebody in South America have been analysed. The data includes copper recovery and the results of three other tests: Bond mill index test; resistance to abrasion and breakage index; and semi-autogenous grinding power index test. A genetic algorithm is used to impute missing data at some locations so as to make use of all 930 samples. The distribution of the variables is modelled with D-vine copula and predictions of copper recovery are compared with those from regressions fitted by ordinary least squares and generalized least squares. The D-vine copula model had the least mean absolute error. IntroductionIn this paper we compare the use of D-vine copula, generalized least squares (GLS), and ordinary least squares (OLS) for modelling geometallurgical data from an orebody in South America. The first objective is to construct models for predicting copper recovery (Rec) from the Bond mill index test (BWi); resistance to abrasion and breakage index (A*b); and semi-autogenous grinding (SAG) power index test (Spi). This involves fitting a D-vine copula and regression models fitted by OLS and GLS. The second objective is to investigate the performance of the fitted models for predicting Rec (Willmott and Matsuura 2005).Traditional resource model approaches either ignore the mineral processing characteristics of extracted tonnages or treat processing as an independent component of a mining operation. The net present value (or any other objective) can be truly optimized only by considering the mining operation as an integrated system in which net value is defined as the end-product that the company sells. This approach requires the resource model to be extended to include all relevant rock properties and processing responses."

Mar 1, 2019