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"Rock falls involving a relatively small amount of material are a leading cause ofinjuries in underground mines in the United States. A contributing factor is unwantedblast damage and over-break. A goal of the United States National Institute forOccupational Safety and Health is to reduce the number of ground fall-related accidents by improving drift driving practices through the introduction of perimeter control blast designs. This paper presents a blast design technique based on a hydrodynamic method. Using commonly available explosive and geometry data, the theoretical particle velocity is calculated as a function of distance away from a cylindrical charge. A calibration factor is introduced to account for actual conditions by comparing the predicted values to a set of field data. The calibrated curve can then be applied for design. The paper introduces modifications to account for different explosives from those used in a calibration test. The application of the approach is demonstrated with respect to a data set from a large-scale concrete block test."

Jan 1, 2009

By N. C. Clarke, T. Wu, M. le Page, K. J. Henley

In 1994, Gold Mines of Australia acquired the rights to the historic Mt. Lyell copper mine, and commenced a feasibility study into expanding production. The ore was shown in laboratory tests to be amenable to single-stage, fully autogenous grinding (FAG) of underground primary crushed ore, and this route offered the maximum cost savings. Flotation performance was expected to be enhanced. However, batch flotation tests done on pilot-scale autogenously ground ore showed an unexplainable recovery loss of between 1% and 3% Cu. A second pilot plant trial was set up, with the ground product feeding banks of continuous flotation cells. Single-stage FAG, FAG/pebble and FAG/ball circuits were tested. Results confirmed a loss of about 3% recovery with single-stage FAG milling. The FAG/ball circuit eliminated the loss. Analyses showed that autogenous milling probably produced more chalcopyrite/pyrite composites and that recovery of these composites was lower than with ball milling. In size fractions below about 53 microns, where chalcopyrite is well liberated, copper recovery was less affected by milling method and flotation selectivity. It was concluded that for the Mt. Lyell situation, a FAG/ball circuit was preferable to single-stage FAG.

Jan 1, 2002

By R. Z. Pyrih

Pressure/ ,acuum lysimeters were used to instrument column tests containing several lifts of tailings slurry, and to sample pore water as the tailings lifts dried. The results of the column testwork demonstrated that air oxidation of sulfide minerals associated wilh the tailings residue neutralized the alkalinity of the tailings pore water. Neutralization of the alkaline pH volatilized cyanide which, in turn, destabilized copper-cyanide complexes and caused the precipitation of insoluble copper from solution. Analyses of pore water samples for microbial activity indicated high plate counts. The presence of viable microorganisms in the pore water implied that degradation of cyanide was enhanced by growth and activity of cyanide-consuming microorganisms in the tailings material.

Jan 1, 1991

By J A. Rabe, PAP Fouche, K O?Neill

Two of the most important issues mines have to deal with are tracking the ore mined from pit to plant and to test the efficiency of beneficiation equipment. Kumba Resources and Saco Systems are working together at moment to develop a RFID tracer system to help address these issues. This paper will focus on the tracking of ore from pit to plant. The main advantages for ore tracking is to ensure plant feed blending and dilution according to plan. The RFID tracer system works on the basis of a tracer that physically simulates the ore properties and radio frequency technology that facilitates communication with the tracer and allowing information to be stored on the tracer. This paper tends to indicate the progress on the application of RFID tracer technology in the minerals processing environment as tested at Kumba Resources so far. It is concluded that the system has great potential, but some important problems remain

Jan 1, 2005

By K. A. Brame, L. G. Austin, R. S. C. Rogers

The breakage parameters of phosphate ore were determined by batch first-ordering grinding studies in a laboratory ball mill according to the Austin, Klimpel, and Luckie procedures. The results were scaled to five larger mills of 0.56, 0.82, 8.36, 4.41, and 5 m in diameter assuming equilibrium ball charges for the full-scale mills. The results were used to predict product size distributions and capacities for batch grinding in the smaller mills and open-circuit grinding in the larger mills. The predicted capacities were in approximate agreement with experimental values except for the 5-m diam mill. The Bond method gave capacity predictions that were 4570 to 150% too high.

Jan 1, 1987

Effective in the 1987 Minerals Yearbook, the following changes have been made: "Asphalt" has been deleted; "Greenland" is included in the "Potash" chapter; "Meerschaum" is included in the "Clays" chapter; and "Staurolite" is included in the "Abrasive Materials" chapter. QUARTZ CRYSTAL Cultured quartz crystal production increased significantly in 198?, following 2 years of reduced production. Consumption of domestic lascas as feedstock for cultured quartz crystal increased. Imports of lascas from Brazil also increased. Although no lascas was mined in the United States during the year, the domestic producer supplied material to consumers from company stockpiles. Specimen-quality natural quartz was very much in demand; quartz crystal for this application is discussed in the Gem Stone chapter.

Jan 1, 1989

By Heath E, van Calsteren P, Turner S, McDermott F

U-Th isotope studies of recent volcanic rocks may be used to evaluate the depths and rates of melt generation beneath mid- ocean ridges and ocean islands. Island arc rocks exhibit a similar range in Th isotopes to that seen in both MORB and OIB, but while those with high (230Th/232Th) tend to have excess 238U, many of those with low (230Th/232Th) are in isotope equilibrium. The excess 238U is attributed to the preferential mobilisation of U in hydrous fluids from the subducted slab within the last 100,000 years, whereas the low Th isotope ratios in the higher-K arcs are primarily due to the introduction of subducted sediment. Preliminary results from mineral separates from lavas on Soufriere on the island of St. Vincent indicate that some of the minerals crystallised up to 50,000 years prior to eruption.

Jan 1, 1995

By C. F. Austin

lined-cavity shaped charge is an explosive mass with a cavity at one end and the detonator at the opposite end. The cavity is lined with a dense material, such as metal, glass, or a ceramic. Such an explosive device can be a useful drilling tool. Lined-cavity shaped charges are constructed to do just one thing — create a high-velocity jet of liner particles that can bore a hole in a target. The liner must be smooth and symmetrical, and high-velocity, high-brisance explosive must be used rather than ordinary commercial dynamites. Since the depth of target penetration is related to jet length, a standoff distance should be provided to hold the charge a short distance from the target. This allows the jet to lengthen prior to target impact. The best results are obtained when the jet is perpendicular to the target surface. Charges can be positioned for firing by taping them to a stake or simply laying them on a flat surface at the proper distance and angle with respect to the target. EFFECTS ON THE TARGET When a jet strikes a rock target, the jet and target particles at the jet-target interface move outward radially with respect to the jet axis to create a tremendous compressive load on the target material immediately below the target surface. The result is a shallow crater formed by spalling at the point of target entry. After cratering ceases, the steadily deepening hole continues to form as a result of the jet's imparting radial motion to the target particles at the jet-target interface. These radially moving particles collide with one another and with the hole walls, the net motion of the spent jet particles and target particles being back out of the hole in an axial direction (the backblast). The backblast keeps the penetration free of debris during penetration, al- though loose surface rock often falls into the hole after penetration ceases. The sudden stress relief after penetration ceases sometimes results in minor hole-wall rebound, which can partially fill the hole. The hole bottom will taper as the jet velocity decreases, until the jet velocity is too low to maintain the backblast. At this point the remaining jet and the following slug material are simply packed into the hole bottom. HAZARDS INHERENT IN USE Air blast from the highly brisant explosives used in shaped charges can damage nearby structures

Jan 1, 1961

By Vladimir P. Nebera, Kostas A. Matis, Piotr M. Solozhenkin, Anastasios I. Zouboulis

"Considered combination of two processes-selective sorptions of metals on biosorbents, inorganic sorbents and their following flotation by dissolved air(DAF). The process of biosorption, using as sorbent suspended non-living biomass (of Streptomyces genus, S. clavulgerus and S. griseus, fungi-Penicillium chrisogenum and Phizopus arrhizus, active slime) in fine particulate dispersion, combined with the application of flotation, for the consequent solid / liquid separation of metal-loaded biomass, has been studied for the removal of toxic and noble metals from dilute aqueous solutions. Biomass of microorganisms possesses great sorptive capacity, exceeding capacity of synthetic and natural sorbents. For instance, Micrococus luteus sorbed gold for 2h Au 253mg/g.The optimal conditions of effective flotation (more than 90%) of actinomyces biomass, loaded with toxic metal-cadmium were found. The adsorpbing flotation technique for copper, arsenic (III, V), chromium removal of metals on pyrite mineral fines (suitable industrial by- product) is an effective process.Dissolved-air flotation was also apllied in the case of dolomite/lead aggregates removal. A fatty acid was used as a flotation surfactant. The presence of dolomite fines aids the efficiency of the process, in comprarison with the flotation of alone lead hydroxide.A laboratory investigation was undertaken on the gold-thiourea system over the whole pH range, aiming at the separation complex of gold [Au(CS(NH2)2]2+ by flotation from dilute solutions. The reagents were found to improve gold recoveries remarkably: the recovery reached over 95%. A separation technique has been investigated, involving gold adsorptionon powdered activated carbon. Foam flotation was employed for the subsequent solid/liquid separation, using sodium oleate as surfactant. The content of Au were 500 mg/L in froth and recovery gold about 100%."

Jan 1, 2003

Single particle breakage tests have been used by a number of researchers to investigate some salient features of the complex comminution process. The characteristics of the common single particle breakage tests such as single impact, double impact and slow compression tests are examined and the results from these tests are compared. A comparison of the results from single particle breakage tests with grindability and ball mill tests is given. Application of the results from single particle breakage tests to modelling industrial comminution processes is described. The necessity for further research into single particle breakage tests to develop a simple but comprehensive technique for estimating the breakage characteristics of ores is discussed.

Jan 1, 1986

By T. A. Marcotte, D. J. Bobrow, S. Vaghar

This paper discusses the geotechnical instrumentation program on the Central Artery/Tunnel Project, in Boston, Massachusetts. Particular attention is given to a state of the art Graphic Information System (GIS) application which has been developed for rapid analyses and reporting of data. This application will see extensive use during the ongoing complex downtown tunnel construction. The excavations in soft ground require cuts of up to 120 feet deep and 250 feet wide. Deformation and groundwater monitoring instruments have been installed to monitor the stability of the excavation support systems and the abutting buildings. The use of this application is illustrated with an example from an ongoing contract.

Jan 1, 1997

By C. K. Eilerts

Finite difference equations were programmed and used to integrate the second-order, second-degree, partial differential equation with variable coefficients that represents the transient linear flow of gas-condensate fluids. Effect was given to the change with pressure of the compressibility factor, the viscosity, and the effective permeability and to change of the absolute permeability with distance. Integrations used as illustrations include recovery of fluid from a reservoir at a constant production rate followed by recovery at a declining rate calculated to maintain a constant pressure at the producing boundary. The time required to attain such a limiting pressure and the fraction of the reserve recovered in that time vary markedly with properties of the fluid represented by the coefficients. Fluid also is returned to the reservoir at a constant rate, until initial formation pressure is attained at the input boundary, and then at a calculated rate that will maintain but not exceed the limiting pressure. The computing programs were used to calculate the results that would be obtained in a series of back-pressure tests made at selected intervals of reservoir depletion. If effect is given to the variations in properties of the fluid that actually occur, then a series of back- pressure curves — one for each stage of reserve depletion — is required to indicate open-flow capacity and related flow characteristics dependably. The isochronal performnce method for determining flow characteristics of a well was simulated by computation. INTRODUCTION The back-pressure test procedurel is based on a derivation of the equation for steady-state radial flow of a gas, the properties of which are of necessity assumed to remain unchanged in applying the test results. The properties of most natural gases being recovered from reservoirs change as the reserve is depleted and pressures decline, and the results of an early back-pressure test may not be dependable for estimating the future delivery capacity of a well. The compressibility factor of a fluid under an initial pressure of 10,000 psia can change 45 per cent and the viscosity can change 70 per cent during the productive life of the reservoir. There are indications that the effective permeability to the flowing fluid can become 50 per cent of the original absolute permeability before enough liquid collects in the structure about a well as pressure declines to permit flow of liquid into the well. The advent of programmed electronic computing made practicable the integration of nonlinear, second-order, partial differentia1 equations2.3 pertaining to flow in reservoirs. Aronofsky and Porter4 represented the compressibility factor and the viscosity by a linear relationship, and integrated the equation for radial flow of gas for pressures up to 1,200 psi. Bruce, Peaceman, Rachford and Rice5 integrated the partial differential equations for both linear and radial unsteady-state flow of ideal gas in porous media. Their published results were a substantial guide in this study of integration of the partial differential equation of linear flow with coefficients of the equation variable. The computing program was developed to treat effective permeability as being both distance-dependent and pressure-dependent. In this study all examples of effective permeability are assumptions designed primarily to aid in developing programs for giving effect to this and other variable coefficients. The accumulation of data for expressing the pressure dependency of the effective permeability is the objective of a concurrent investigation. FUNDAMENTAL DIFFERENTIAL EQUATIONS DIFFERENTIAL EQUATION OF FLOW Linear transient fluid flow is represented by the partial differential equation,

Jan 1, 1965

By Bahaa Hussein

Considerable number of methods and procedures has been introduced for trending and progress reporting of engineering, procurement and construction projects. This paper introduces a novel method, designed to augment the set of indices for measuring schedule performance. The concept behind the developed method is to provide early warning by dynamically evaluating the schedule cumulative impact of construction projects. The proposed method is designed to highlight the gradual deterioration in schedule status resulting in a cumulative impact due to the aggregation of small delays encountered during the execution of non-critical activities, and can be used as an add-on utility to existing software systems that perform CPM scheduling. The proposed method is based on a developed formulation for quantification of the Schedule Compression Index, which takes into consideration the remaining project activities durations. The index is developed based on periodic comparisons of ?Current Schedules? and the project ?Baseline Schedule? using the summation of the remaining activities durations. A numerical example is presented to demonstrate the application and capabilities of the developed method. It is expected that the developed index together with other schedule performance indices would be of value to decision makers and members of project teams.

Aug 1, 2013

By R. E. Gilmore

This paper is a revision of a former paper published in mimeograph form by both the Canadian and American coal classification committees, and is now presented for the purpose of acquainting those interested with the present status of coal classification in Canada. The classification employed for "Customs purposes" is detailed, and the classification schemes from which the customs classification was derived are reviewed and discussed. The activities of the Canadian coal classification committee, namely, the Associate Committee of Coal Classification and Analysis of the National Research Council, is also reviewed. Reserves, Production and Uses of Canadian Coals The total coal resources of North America are estimated to be 69 per cent. of those of the world, and of this portion, the distribution is 52 per cent. for the United States and 17 per cent. for Canada. Accordingly, Canadian coal resources amount to approximately one-third of those of the United States, and one-sixth of those of the world. As for the geographical distribution, the bulk of the known coal deposits are in the extreme eastern and western portions of the country. Employing the usual, more or less loosely defined, designations for the different classes and subclasses of coal, the distribution of the Canadian coal resources, according to provinces, is as follows: Maritime Provinces: Nova Scotia and New Brunswick Bituminous and semibitu- minous coals. Ontario (northern part).............................Lignite. Saskatchewan, (and Manitoba)....................... Lignite. Alberta and British Columbia........................ Lignites, subbituminous, all kinds of bituminous, semibituminous and semianthracite coals. Of the total reserves, roughly 92.5 per cent. is credited to the noncok-ing lower rank subbituninous and lignite groups, the bulk of which are

Jan 1, 1930

By W Frint

An area in Southwestern Wyoming is rapidly becoming the soda ash capitol of the world. Starting only 20 years ago from a unique deposit of trona, the out-put has rapidly increased until now about half of the soda ash produced in the United States comes from this area. Chemically, soda ash is sodium carbonate, a large tonnage, widely used alkali. The largest single industrial use is in glass manufacture, where 25 tons of soda ash are used in producing 100 tons of glass. Other big users are the detergent and chemical industries. The trona deposits in the Green River basin of Wyoming, were formed about 75 million years ago as part of a geologic period when an ancient lake alternately went through periods of being a large fresh water or smaller brine lake. Today these deposits lie from 800 to 3000 feet below the surface in a semi-arid, almost desert region just west of the Rocky Mountain Continental Divide at an elevation of 6200 feet. In this area there is a prevailing westerly breeze and almost every afternoon the velocity reaches very windy conditions. Rainfall is minimal, being only 6-8 inches of moisture per year with an evaporation rate exceeding 24-inches. Most of the moisture comes in winter snows or summer thunderstorms. Temperatures are low with below zero weather dur¬ing most of the winter and early morning frost is not uncommon during the summer months. There are many trona beds varying in thickness from a few inches up to 15 feet. The deposits now being mined vary from 8 to 10 feet in thickness. The ore is mined by driving tunnels either via conventional explosives or with continuous boring machines. It is transported to the shafts via conveyor belts

Jan 1, 1972

By Clayton V. Deutsch

Indicator Kriging (IK) is being commonly applied to mineral resource assessment problems. More recently, indicator simulation has also been used for sensitivity analysis and difficult change of support problems. The essential theory underlying indicator geostatistics has not changed since its introduction 10 years ago; however, the last decade has produced many important developments in its implementation. The objective of this paper is to consider some of the more important implementation details. Examples from the Porgera gold deposit in Papua New Guinea are used to illustrate the steps in an indicator geostatistical study and the importance of various implementation decisions. Specifically, in many precious metals deposits the continuity of the grades is typically quite anisotropic. This anisotropy varies in both magnitude and direction throughout the deposit. Building a model of this anisotropy can considerably improve the local kriged or simulated points. One approach to build this model is presented. In addition, there is typically more than one mineralization style and geological rock type; the spatial character of the mineralization can be quite different within each geological rock type. The rational treatment of anisotropy and geological subdivisions is common to all mineral resource assessment procedures; the importance of this qualitative geological information is not diminished with more sophisticated techniques such as indicator kriging. This paper presents an overdue comprehensive look at the practical aspects of indicator methodology. Many of the techniques and procedures documented in this paper have been developed during the course of detailed mine feasibility studies.

Jan 1, 1992

By Shuji Owada, Mayumi Ito, Idemitsu Kosan Co. Ltd., Takehiko Nishimu, Sadayuki Shinozaki, Takashi Ando, Tomoe Ota

"fElectrical disintegration (ED) by high voltage electrical pulse is a relatively new technique which was demonstrated by Tomsk and Mekhanobr in the 1960’s and has been investigated in detail by Andres in the 1990’s. In the ED, the electrodes should contact to the targeted material and the energy of high voltage pulse is charged directly to the material. It has been reported that the ED is quite efficient to improve liberation of composite materials, because electrical current passes often through the conductive part and/or along the phase boundaries. Theoretically, the energy required in the ED should be lower than in the other mechanical breakage methods, because breakage occurs basically by tensile strength in the ED while in the other mechanical methods it does mainly by compressive strength which is generally ten times or higher than tensile strength.Coal, involving many kinds of mineral matters, is considered to be one of the targets for the ED to apply to. In Japan, national “Hyper-Coal Project” is now running, in which ultra-clean coal (less than 200ppm of ash, 0.5ppm of Na and K, and 2ppm of Ca) is planned to produce by combining advanced coal preparation and liquefaction technologies for the purpose of direct injection of coal combustion gas to turbine. The ED is one of the major techniques to be developed in this project, since mineral matters must be liberated before introducing to the following advanced separation processes. Liberation is important also for the utilization of middling coals generated from usual coal preparation plants. The application of the ED to the middling coals is considered effective for upgrading these coals with higher liberation degree and less energy input."

Jan 1, 2003

By D. E. Kirby

In line with the Bureau of Mines' goal of providing technology to maintain a supply of minerals adequate to meet national economic and strategic needs, a method for recovering chromite from lateritic materials was investigated. Laterites are a large resource of Ni, Co, and Cr, critical metals of which the United States has insufficient domestic sources. A roast-leach method was developed by the Bureau to extract nickel and cobalt from domestic laterites. A concurrent effort was made to recover chromium from either unprocessed laterites or the leach residues. This report characterizes laterites and laterite leach residues. It describes a chromite recovery procedure using conventional industrial beneficiation equipment for sizing, low-intensity magnetic separation, and gravity concentration. Concentrates containing 20.0 to 22.4 pct Cr (29.0 to 32.7 pct Cr203) and with a 1.6:1 to 1.9:1 Cr-Fe ratio were recovered from laterites and residues containing 1.3 to 2.5 pct Cr with recoveries exceeding 50 pct. Because of their low grade and fine size, these concentrates would be suitable for preparing chromium chemicals. The direct beneficiation of laterites for chromite recovery resulted in the minerals containing nickel and cobalt concentrating in slime products. No iron oxide coproducts of acceptable grade were recovered from either laterites or residues.

Jan 1, 1982

By B. S. Crocker

Introduction Pebble mill grinding can be defined as wet or dry grinding in a tumbling mill with pebbles used as grinding media instead of metal balls. To clarify the terminology, which may differ in the various mineral industries, it is essential to point out the differences between grinding with partially sized and carefully sized grinding media. When grinding with partially sized media (limited top size only), sizing is usually the setting of the primary crusher and there is no bottom sizing-all fines, chips, sand, and small sizes are included. This is often called run-of-mine or primary autogenous grinding. In this case the entire crusher discharge is fed into the mill and part of it becomes "charge" and the rest is "sand feed." When grinding with carefully sized media (i.e., media that is sized both at the top and bottom of the size range), no fines or chips should be present in the charge to the mill. This is pebble milling and the grinding media consists of ore, rock, Danish pebbles, gravel pebbles, or ceramic pebbles-media that is appreciably lighter than steel balls. The grind¬ing charge is fed to the mill to maintain a load and it does the grinding usually on a classified sand product of suitable size. If the grinding media is the same as the raw material being ground, this type of grinding is also called secondary autogenous grinding. Pebble milling was in common use in the gold mines of South Africa in the early 1900s. It is by far the oldest form of autogenous grinding. It may be used to grind crushed ores from a maximum size of about -% in. to all -10 µm. Pebble milling differs from "run of mine" grinding in many ways, but two big differences are: I) The ore itself is crushed to approximately t/4 in., or it may be further reduced to 8 or 10 mesh in a rod mill or fine crusher. 2) Then carefully screened ore of the correct size is placed in one or more stages of pebble mills to grind the ore to its desired final size. The media size is carefully selected to be the most efficient for the particular size reduction required in each stage. This is a precise practice, and pebble mills can be designed as easily and accu¬rately as can rod mills or ball mills by those skilled in the art. A pebble mill is a rotating cylinder in which the grinding media has been carefully sized for the job in hand. If the mill is doing coarse grinding, the media will be correspondingly large and the mill might be called a primary (or intermediate) pebble mill or a lump mill. If it is to do fine grinding, the media selected is proportionately smaller and the mill would be called a secondary or tertiary pebble mill. The media itself is normally rounded pebbles made from screened ore from the primary-crusher discharge. The charge could be rounded pebbles of the correct size removed from a primary autogenous mill, or the pebbles could be carefully sized gravel, sized beach pebbles, or sized flint pebbles. Usually the specific gravity of a pebble is from 35 to 55% of that of steel balls. It commonly varies from 2.66 to 4.2, depending on the specific gravity of the ore. Primary pebble mills, or lump mills, would be charged with pieces of rock up to 10 in. maximum and 5 in. minimum and would grind crusher discharge with a maximum size of 1/4 in. on hard ores and 3/8 in. on softer ores. Secondary pebble mills could be charged with pieces of rock screened on 4 x 4-in. square mesh and retained on 1 1/2 x 1 1/2 in. These mills would grind material with a top size of 8 to 10 mesh. Tertiary pebble mills could be charged with screened rock that would pass 3 x 3 in. and be retained on 1 x 1 in. and grind 35 to 48 mesh sands. Principles of Operation Media Size To operate a pebble mill efficiently it is imperative that great care be taken in screening the rock feed that is fed into the mill to round up into a range of pebble sizes that are suitable for grinding of the sands that are also being fed into the mill, having regard to the top size and the average size of such sands. Since most mill operators know from experience the correct size of steel balls to be used in a mill to handle different sizes of sand feed, it is convenient to relate the pebble size to a ball size. If the correct ball size is not known, it is relatively simple to conduct batch laboratory tests in 12- to 18-in. laboratory mills to determine the optimum size of steel ball. Also, Bond and others have published formulas for calculating the optimum ball size. When the correct size of the individual steel ball is known, then a pebble of the same weight will perform the same grinding if used in the proper size of mill and under the proper conditions, which will be defined later. This media size, as defined by the average weight of the individual pieces, is most important. The following tests will serve to clarify this point. Laboratory Tests Using Steel Balls or Pebbles. Several labora¬tory grinds were made to compare a given size and weight of steel ball with a pebble of the same weight but larger size, and of the same size but lighter weight. The following loads were used in the 12-in. laboratory mill: All 1 1/4-in. steel balls weighing 122 ± 10 g. All 1 3/4-in. pebbles weighing 123 ± 5 g. All 1 1/4-in. pebbles weighing 45 ± 3 g. To equalize capacity, the test with the steel balls was run only 34% as long as the tests with pebbles. In each case, 2,500 g of primary sands were ground at the optimum gravity for the media being tested at a speed 78 1/2 % of critical speed, and for 170, 340, and 680 revolu¬tions for the steel balls, and 500, 1,000, and 2,000 revolutions for the pebbles. The first test is summarized in Table 74 by showing the 80% passing size in microns for the feed and for each of the grinds. Two other tests are summarized in Table 75 showing the screen analyses

Jan 1, 1985

By R. D. Brown

The first step in achieving a rationalization of Canada's crazy patch-work quilt of mining taxes and policies was achieved with the release last November of a major federal-provincial report on the taxation of the Canadian mining industry. The Report, prepared by a group of federal and provincial civil servants, sets out a concise and informative discussion of the major taxation and incentive issues facing the mining industry, and puts forward major recommendations for more rational and moderate taxation policies. It will be considered by federal and provincial Finance and Resource Ministers for possible action by the respective governments. The Report was prepared in response to a directive from the federal and provincial First Ministers in February 1978 to review the federal and provincial taxation of the resource areas. Although the Report is more successful in identifying problems rather than concrete solutions, and reflects some continuing differences of opinion between the governments involved, it does represent an important first step in a comprehensive review of our resource tax policies. This article summarizes the Report, which is some sixty pages in length. Although a detailed discussion of it is beyond the scope of this article, the report itself is recommended reading for those concerned about the current state of the Canadian mining industry.

Jan 1, 1979