Search Documents

Statistics on the imports of industrial minerals into the EECshow that the traded volume in 1987 is approximately 23 million tonnes.The following table documents the imported quantities on the most important industrial minerals.

Jan 1, 1989

By Howard N. Eavenson

WHEN the A.I.M.E. was formed 75 years ago the bituminous coal industry was in its swaddling clothes, although it had been operating for more than a century and coal was being mined in every state now producing it, and its output had long passed that of anthracite. In 18/1, the value of bituminous coal was about seventy per cent of the value of all metals produced and more than one fifth of the value of all products of mines, both metallic and nonmetallic.[1a] While many of the operating mines then shipped by rail, most of the development was along waterways, although the canal era of transportation was nearly at its end. Coal mining is a heavy industry and one of the best measures of its progress is its output throughout the period. Fig. 1 shows this graphically from 1800 to 1945. From 1870 to 1910, the production-which is very nearly the consumption-almost doubled every decade, a most amazing growth. In 1918, in World War I, a peak was reached and a decline began, which ended in 1932. A gradual recovery then started, interrupted in 1938 but culminating in the record output of 1944, when, owing to World War II, a tonnage of 619,576,000 was produced. It is evident that the forces leading to the continuous increase ending in 1918 have spent themselves and that only unusual conditions will require so large a production again. The postwar conditions of World War II will probably

Jan 1, 1947

By Yi-Hsuan Chen, Jürgen Brune, Candace S. J. Tsai, Alejandro Munoz, Connor Krause

This study highlights the significant presence of submicrometer- sized airborne particles in underground coal mines, especially around the belt conveyor drift areas. Using realtime instruments and gravimetric samplers, the assessment revealed that particle concentrations were notably higher in the belt entry compared to other locations, with a strong presence of submicrometer- and nanoparticle-sized dust. The study’s findings emphasize the importance of addressing small particle exposure, as these particles, composed of harmful substances like calcium, copper and iron, can pose serious health risks to miners. This research emphasizes the importance of improving air exposure estimation through more effective monitoring and characterization techniques to better understand and reduce workers’ exposure to hazardous airborne particles in the mining environments.

Apr 1, 2025

By P. Hajiani

"It has been shown that the economics of the hydrometallurgy of a rare earth process depend upon the amount of acid needed to extract the metal per unit mass of the ore. In this paper, we show how to save a significant amount of acid treating a low grade ferrocarbonatite and extracting +99% constituent REEs. In order to save the acid, the alkali earth metals, originated from gangue minerals were removed prior to REE leaching. Moreover, the operating conditions of leaching reactor are optimized to leave iron behind and to leach REE selectively and effectively. Consequently, the hydrochloric acid consumption per ton of ore drops significantly compared to other processes.INTRODUCTIONThe viability of an extractive metallurgy process of rare earth elements (REEs) to produce an intermediate mixed concentrate depends upon a low-cost simple flowsheet (Bogner, 2014). It has been known that regardless of the mineral, the acid consumption is intimately associated with the operational cost (OPEX) of any REE project, particularly when the initial REE grade is low. Besides acid, the other reagents consumed in the process impact the OPEX. This makes minimizing acid consumption in an REE method a critical priority. The acid consumption in different projects has been reported to be equivalent of 100 kg to 1000 kg HCl (100%) per tonne of ore (Average: 500 kg/t HCl 100%) (Bogner, 2014). This value is strongly correlated with the grade of REE in the ore and therefore in the pre-concentrate.Previously, it has been stated that the Montviel flotation concentrate requires about 300 kg HCl (100%)/t ore to attain +99% REE recoveries using direct leaching (Yu, Verbaan, Pearse and Britt 2013). Hence, for an ore feed rate of 2500 t/d ROM, a large acid production plant (744 t/d HCl 100%) would be required for acid leaching of such concentrate. The equivalent Chlor-alkali plant is estimated to cost M$ 150 (CAPEX) and may require 100 MW power. Therefore, there is a necessity to improve the extractive hydrometallurgy and to reduce the acid consumption significantly."

Jan 1, 2016

By Gray N. B, Grimsey E. J

A steady state heat and mass balance simulation of the Western Mining Kalgoorlie Nickel Smelter (KNS) flowsheet was developed within the framework of the commercial software METSIM. The KNS simulation can achieve all plant setpoints through the adjustment of controlled process inputs, and provides a useful tool for the evaluation of the current flowsheet and potential flowsheet developments The simulation model and a number of applications are discussed in this paper.

Jan 1, 1991

By Stefan Basista, Janusz Tomala

Modern constructions of electric ferroalloy furnaces are still based on applying in the lower "liquid" areas the standard carbon materials. What can be illustrated by the experience from several years, in the part of tap hole, hearth walls and bottom, is that so-called carbon or semigraphite microporous materials are successfully being applied. These materials, compared to the standard carbon, give better resistance to the attack by liquid metal and other wear factors. Generally this group includes materials characterised by considerable share of pores of diameters below1µm (significantly exceeding 90% of total porosity). The microporosity increases the resistance to the infiltration of the liquid metal. The effect of microporosity can be obtained by introducing silicon (or its compound) additive to materials based on electrically (or gas) calcined anthracites, electrographite and binder. Moreover, by introducing other ceramic additives, the microporous materials are more resistant to dissolution by the liquid metal. Observing the trends of "improving the materials" in the last years, it is clearly seen that the share of non-carbon additives is bigger and bigger, so we can reasonably speak of rather carbon-ceramic materials. In this paper, the survey of the most important quality criteria, which must be met by the carbon furnace linings, as well as the new testing methods will be presented.

Jan 1, 2007

By André Piché

"Over the last ten years, several developments have been undertaken to conceive a machine capable of carrying out simultaneously all the mining tasks in hard rock deposits: cutting, mucking, hauling, ground support and dust control.The main advantages of continuous mining over conventional drill and blast is the elimination of the cyclical nature of the actual methods. Other advantages sought are improved stability of mine openings, higher productivity, lower costs, and a cleaner and safer mining environment. Noranda has con-tributed to these developments in a major way. As a member of HDRK, Noranda participated in the development of two mechanical continuous mining machines; the TM60 and the Wirth. Noranda also undertook a development of its own based on the utilization of electricity to break rock. This technology is called plasma blasting (PBT). The concept of plasma blasting is illustrated in Figure 1.When the operating voltage is reached (18kV), the energy is released through a high current switch (500k A) into a coaxial powercable. It is then delivered to the tip of an electrode inserted into a drill hole which is filled with an electrolyte (water). When the energy reaches the electrolyte, it is instantaneously transformed into a high temperature, high pres-sure plasma that produces a shock wave capable of breaking rock(1).After initial testing of the plasma blasting technology in the lab and in the feld, two experimental platforms were constructed. The first one, a test bench, was erected at the Noranda Technology Centre in Pointe Claire, Quebec.(2)The second one is a rock breaker prototype based on a standard diesel-powered carrier (Fig. 2). Each platform was equipped with a 300 kJ pulsed-power module.In 1993, the rock breaker prototype was used for an extensive series of tests in the field. During that campaign, over 700 oversized boulders were blasted. The results of the tests demonstrated that more durable components for PBT were needed. The test bench was then used to successfully develop and test long-lasting components (probe, switch, cable, connector) that can be adapted to industrial platforms leading to the commercialization of the technology."

Jan 1, 1996

By George Holderer

THE relation between man-hours of labor and production may be correlated for any industry, and already it has been widely used in piece-work studies. It is not in general use as yet for recording labor efficiency of an enterprise or an industry. An exception to this is the Cement Institute, which tabulates the man-hours used in each operation and FIG. 1.-LABOR PRODUCTIVITY IN COPPER MINING. thus gets the total man-hours per barrel of cement. The writer knows of one steel company that keeps a complete record of man-hours for every operation from mining the ore to finishing the product. No doubt many others find the information useful, and such a study could be of value to the mineral industries.

Jan 1, 1935

By Richard J. Fantel

The U.S. Bureau of Mines performed engineering and economic analyses on 130 domestic phosphate mines and deposits, all containing demonstrated resources, of which 73 are in the Southeastern states of Florida and North Carolina, to determine the domestic availability of phosphate rock. Of the total domestic resources studied (6.4 billion tons of recoverable rock) over 60 percent is found in these two Southeastern states. At an estimated cost of $30 per ton, in January 1981 dollars, approximately 1 billion tons of rock is potentially recoverable from Southeastern deposits, increasing to 3 billion tons at $45. In addition, there is nearly 6 billion tons of recoverable rock at the inferred level (some presently unmineable) in Florida and Georgia.

Jan 1, 1983

By S C. Bhowmik, R Gupta

This paper presents the ventilation problem faced by Jaduguda uranium mine, operated up to a depth of 905 m. The first two stages of the mine workings are up to 555 m, which are nearly exhausted. The third stage of the mine is from 555 m to 905 m and presently most of the stopping and development activities of the mine are carried out in the third stage. The total available intake quantity in the third stage was only 21 m3/s. As a result, the mine was facing a radiation hazard. The different instruments and the methods used for monitoring of radiation levels are also presented in the paper. It was observed that the average individual effective dose (external + internal) of radiation workers in Jaduguda mine increased from 6.93 mSv/y in 1995 to 9.5 mSv/y in 2001 and during this period the average individual effective dose of mine workers of the world was less than 5 mSv/y. Keeping this problem in view, a detailed ventilation investigation has been carried out in this mine. Initially the requirements of intake air quantity have been assessed by carrying out field investigation in different stopes and it was found that the third stage workings require nearly 100 m3/s. A detailed ventilation survey has been carried out in the total ventilation network and a basic ventilation model of the mine has been prepared. Different alternative schemes for effectively ventilating the third stage workings of the mine were simulated and a set of recommendations were made for implementation. The recommendations were partially implemented in 2002. As a result the total intake quantity to third stage workings increased to nearly 70 m3/s and the average individual effective dose (internal + external) of mine workers came down from 9.5 mSv/y in 2001 to 6.65 mSv/y and 6.60 mSv/y in 2002 and 2003 respectively. The implementation of remaining recommendations is in progress. Finally, a mathematical model has been formulated from the first principle for predicting air quantity requirements in cut and fill stopes in uranium mines and this model is being validated for further use.

Jan 1, 2005

By Svenson D

Most exploratory drilling in the N.S.W. coal fields has been conducted using convent- ional diamond drill rigs of modern design, us- ually truck or trailer mounted. Conventional diamond drill rigs have also been used exten- sively by the Queensland Department of Mines for coal exploration. However almost all of the exploratory drilling of surface mine coal prospects throughout Queensland has been conducted using rotary rigs typified by the Mayhew 1000 truck mounted unit. The rotary rig has proved capable of both rapid non-core drilling and core drilling with high core recoveries when modern core barrels have been used by experienced drillers. For deep or continuous core drilling, wire-line equipment is now in general use. Both wire-line and conventional coring is effected using stationary split inner tube barrels, for minimum core disturbance, maximum core recovery, ease and accuracy of logging and sub-sectional sampling. Diamond core bits are generally required in N.S.W., with water circu- lation. In Queensland, the generally softer coal measures strata can be successfully cored using tungsten carbide bits and air circulation is the preferred method where possible.

Jan 1, 1977

By Margaret V. Doyle, R. W. Powers

INTERNAL friction measurements, carried out as functions of temperature, have been used extensively to obtain data on the mobility of interstitial impurities in the Group V metals, vanadium, colum-bium, and tantalum. For a given frequency of vibration, the internal friction peaks at a temperature which depends principally on the particular metal and interstitial impurity. Relaxation times and corresponding activation energies are determined from a knowledge of such peak temperatures for various vibration frequencies. These experimentally determined relaxation times are related to the average jump frequencies of the interstitial impurity atoms using the model of Snoek1 and can be related to macroscopic diffusion coefficients as shown by Polder.2 In these Group V elements, the impurity whose diffusion rate has been the hardest to pin down by the internal friction technique has been carbon. The experimental difficulties associated with carbon in these metals were not evident in the earliest investigation carried out by Ke on tantalum.3 The carbon peak was thought to occur near 140°C at 1 cps on the basis of the enhancement of the height of this peak measured on specimens which had been coated with colloidal carbon and heated in vacuo at 1200°C. Presumably on the basis of similar evidence, Wert identified a peak also near 140°C at 1 cps in columbium with the diffusion of carbon in that element." Later, the difficulties of admitting carbon into these metals without simultaneously introducing oxygen were pointed out by Ang and Wert- nd by Marx, Baker, and Sivertsen.6 The latter suggested that the peaks attributed to carbon in tantalum and columbium might arise from the presence of traces of oxygen, adventitiously absorbed during the carburization. Such suspicion was confirmed, at least in tantalum, when it was found possible to prepare C-Ta specimens without contamination by oxygen. Specimens sci prepared, aitnough high in carbon content, gave no evidence of a damping peak near 140°C.7 Recently, the authors were able to Show that carbon diffuses in tantalum at a rate only slightly slower than that for nitrogen.8 At 1 cps, the carbon peak lies at 34g °C, only a few degrees above the nitrogen peak at 341 °C. However, the two could be distinguished by a different stability with respect to aging at 400 °C and by a different dependence of peak height on solute concentration. On the basis of this experience with tantalum, it appeared worthwhile to search also for a carbon peak in columbium in the vicinity of the nitrogen peak." In this report, the location of the carbon peak in columbium is made known and its position relative to the nitrogen peak is compared with the authors' previous findings in tantalum. In addition, a second source of internal friction seen only in specimens of very high purity is discussed. Columbium was obtained from the Fansteel Metallurgical Corp. as either 30 or 40 mil diam stock. The manufacturer gives the following analysis as typical of this material: 99.4 pct minimum Cb, 0.10 pct maximum C, 0.5 pct Ta, 0.04 pct Ti, 0.02 pct maximum Si, and 0.01 pct maximum Fe. Wires 13 in. long were degassed near 2200°C for 4 to 16 hr in an ultimate vacuum better than 2.10" mm Hg in an apparatus described previously.8 This treatment reduced the oxygen and nitrogen damping peaks observed in the as-received material to less than 0.0001 in Q-1 above background. The latter value varied somewhat from specimen to specimen in the range 0.0004 to 0.0008, depending chiefly on the wire diameter and the frequency of vibration. The high temperature evacuation left the specimen in a rather soft condition, the Vhn being reduced to 40 from a value near 121 in the as-received conGi~iur~. The gram size was about the diameter of the specimen. After outgassing, specimens were loaded with carbon by absorption of such hydracarbons as methane or ethylene. Carbon was also added to a few speclrnens by heating them at

Jan 1, 1958

By M. H. B. Nasseri

Barre granite?s samples have been tested for fracture toughness under mode I along two different directions with respect to microstructural fabrics were investigated. 3D objects counter algorithm technique is applied on micro CT images to identify physical properties such as crack induced porosity, crack density, generated total surface area and contribution of individual mineral grains within the damaged zone.The result obtained justifies very well the fracture toughness anisotropy in Barre granite, and explains the difference in the structure of the damage zone evolving as a result of test crack interacting with pre-existing microstructural fabrics. The present study has potential application in the fields of rock mass stability, oil reservoir stimulation, rock fragmentation and damage in blasting and emplacement of dykes.

May 1, 2009

By A. Dadgar

The adsorption kinetics of AuBr4 on Dowex ® -21K ion-exchange resin were studied and found to be independent of the initial gold, bromine, and base metals concentrations. A commercially-available undivided hypochlorite generating cell was used to generate bromine from simulated mixed halide (Br-/CI-) solutions. Electrolyses were performed with solutions containing very low Br' ion concentrations. The current efficiencies were 80-90%, and Br- ion utilization was maximized. A detailed economic assessment was made considering the capital and the operating costs for a 1000 mtpd gold ore processing plant. The toxicity and discharge levels of bromide, bromine, and cyanide were compared from a detailed analysis of the published literature data. Comparing bromine vapor with hydrogen cyanide gas for a given length of exposure, the lethal concentration of bromine is about 3 times less than hydrogen cyanide. The toxicity of sodium bromide is similar to that of sodium chloride. Acute poisoning to bromides and chlorides is rare because of their low toxicities. The allowed discharge levels of Br-, Br2, and CN- in fresh water are 250, 0.1, and 0.005 mg dm-3, respectively.

Jan 1, 1992

Reduce workers' dust exposure by capturing airborne coal and quartz dust generated by a continuous mining machine. Approach Install a water-powered scrubber system operating at 1,000 psi on the boom of a continuous miner at the conveyor throat. Dusty air from the face area is drawn through the scrubber and cleaned air discharged into the return. How it Works Most coal mines use exhaust face ventilation to send fresh air over the miner operator and contaminated air behind the line curtain to the return. The problem with exhaust ventilation is that the line curtain should be within 10 ft of the face for maximum effectiveness. If the line curtain is not extended properly, or water sprays for dust control are poorly installed, dust rolls back towards the operator. An effective way of removing this dust is with a scrubber mounted on the mining machine.

Jan 1, 1988

By Graham W. Marks, Clyde A. Bridger

"INTRODUCTION An electrolytic process for the preparation of metallic cobalt has been developed recently in the Bureau of Mines Experimental Station at Boulder City, Nev. Chemical analysis shows this cobalt to contain impurities in small percentages, chief of which is nickel. It is desirable to be able to determine these impurities spectrographically. This report is concerned with a procedure for the spectrochemical determination of nickel in cobalt.Previous spectrographic work on cobalt was done by Breckpot and Mevis,/ who used the method of homologous pairs of lines. However, it was felt that a more detailed investigation was needed, employing the method of internal standards.ACKNOWLEDGMENTWe are indebted to B. G. Knickerbockor and his staff at the Boulder City, Nev., station of the Federal Bureau of Mines for their cooperation, and particularly to Richard W. Lewis and Robert G. Sanderson, who prepared nickel-free electrolytic cobalt for us."

Feb 1, 1948

By Forrest A. Kerr

Northern British Columbia is a term which probably conveys to various individuals quite different meanings. To the average person it signifies an area of indefinite size and indefinite position which may start at Prince Rupert, Stewart, or somewhere else, and end at Alaska or .the Yukon. Many people, in their conception, fail to encompass in their mind's eye that vast tract of land lying ~o the east of the panhandle of Alaska. Few realize that the area north of an east-west line through Stewart represents more than one-third of the Province and is, of course, the part most appropriately designated as northern British Columbia. This section, more especially the northernmost 100,000 square miles, from political, geographic, and economic viewpoints stands isolated perhaps more than any other part of British Columbia. Unlike the rest of the Province, it has no coast line, being cut off from the Pacific ocean by a narrow strip of Alaska. It is bounded on the north by the Yukon, and on the east by Alberta. In the south this northernmost 100,000 square miles is limited by a divide of rugged mountainous country separating the drainage basins of the Stikine and Liard rivers from those to the south.

Jan 1, 1929

By Max J. Spendlove

The secondary copper industry is comprised of numerous enterprises which collectively employ many of the recovery and refining processes now used in primary plants as well as many other processes that are unique to the industry. Effective methods are used for identifying and segregating all types of scrap according to widely accepted standard classifications. Segregated scrap and waste materials usually require some preliminary processing to remove both valuable and deleterious associated constituents. Smelting, melting, alloying, and pyrorefining are common methods used for producing secondary metal, but operating techniques usually differ from primary metal operations because of the difference in physical and chemical properties of the respective raw materials. Some copper scrap is converted to copper chemicals rather than to refined metals or alloys. Recovery efficiencies of all processes are reasonably high, and the quality of products meets rigid specifications.

Jan 1, 1961

By R. C. McFarlane, T. D. Mueller, J. E. Walstrom

In evaluating uncertainty, experiments are usually performed repeatedly and then conclusions are drawn from the distribution of results. With the advent of high-speed electronic computers, it is possible to perform experiments using mathematical models constructed to simulate complex experiments or operations. Statistical methods are then applied to the results of the simulated experiments. This procedure forms the busis of this paper. Demonstrated is the need for properly accounting for uncertainty in petroleum engineering problems. How uncertainty affects solutions is evaluated in three example illustrations. The method used to evaluate uncertainty in petroleum engineering studies is the Monte Carlo simulation procedure.'-" INTRODUCTION The solution to most technical problems may be derived from interrelationships among several quantities called variables or parameters. There may be only a few variables or several hundred. Interrelationships among parameters may be explicit or implicit, well established or only approximate. Some variables that fully or partially depend on the magnitude of others are called dependent variables. Input variables for most practical problems are not precisely known; there is usually an uncertainty in their value. The degree of uncertainty may vary from one variable to another. Variables that are known accurately are called determinates.' For instance, the gravity of crude obtained from a particular pool may be known precisely, and therefore is a determinate. The degree of precision with which a quantity can be determined increases as data describing the pool are accumulated during the development of the field and the producing life of the pool. The uncertainty of a parameter may result from difficulty in directly and accurately measuring the quantity. This is particularly true of the physical reservoir parameters which, at best, can only be sampled at various points, and which are subject to errors caused by presence of the borehole and borehole fluid or by changes that occur during the transfer of rock and its fluids to laboratory temperature and pressure conditions. Uncertainty may also result in attempting to predict future parameter values. This type of uncertainty is particularly evident in investment analyses involving future costs, prices, sales volumes and product demand. Uncertainty in the solution to investment problems is often called risk, and its study is called risk analysis.' Uncertainty also enters into biological and sociological analyses in which indeterminate factors are often important due to limited control of the experimental material. It is customary, in evaluating uncertainty, to perform repeated experiments and to draw conclusions from the distribution of the results of these experiments. With the advent of the high-speed electronic computer, it is possible to construct mathematical models which simulate complex experiments or operations and to perform the experiments repeatedly, utilizing the models. Statistical methods are then applied to the results of the simulated experiments This method forms the basis of the investigation reported here. PROBABILITY DISTRIBUTIONS FOR VARIABLES The uncertainty in the value of a variable may be indicated by a probabilistic description accomplished by expressing the quantity by a probability distribution. Many recognized probability distributions can be used to describe physical quantities. Recent studies used various types of distributions to describe core analysis data.',' However, for the examples in this paper, the uniform and triangular distributions are believed to reasonably approximate the data used (Fig. 1). The uniform distribution confines the variable between an upper and a lower limit. The variable may lie anywhere between the two limits. This distribution is used when no one range of values for a variable is more probable than any other, but information or intuitive reasoning indicates the variable will lie somewhere between the chosen limits. The triangular distribution is used for a variable when more data are available to indicate a central tendency of distribution. This allows postulating a "most likely" value to the distribution and upper and lower limits. In this case, as for the uniform distribution, the variable is not expected to assume a value less than the lower limit or greater than the upper limit. However, with improved quality of data it can be postulated that the variable will tend to assume a value close to the most likely value, and that there will be a decreasing probability for values away from the most likely value. The area under either of these probability distributions is equal to unity since it is assumed that there is a 100 percent probability that the variable will lie somewhere under the curve. An ordinate erected at any particular value of the variable divides the area under the curve into two parts: the area to the left of the ordinate represents the probability that the value of the variable will be equal to or less than the value of the variable at the position of the ordinate, and vice versa. The probability is zero that the variable will have any specific deterministic value. If two ordinates are drawn for any two values of the variable, the probability that the variables will have a value lying between these ordinates is equal to the area under the curve lying between the ordinates.

By W. A. Crosby, A. Bauer

The recent increase in size of open-pit mining operations has resulted in major improvements in efficiency that are beneficial to the mining industry. Higher bench heights, larger diameter blast-holes, and more powerful explosives have played an important part in reducing mining costs. These measures have, however, resulted in increased energy concentrations in the blast area, which can result in severe backbreak problems for final pit walls. If backbreak is not controlled, a decrease in the overall pit slope angle will ultimately be necessary, and consequences, such as decreased recoverable ore reserves and increased waste-to-ore ratios, will result. Greater amounts of loose face rock will be produced, and planned safety berms will be less effective or nonexistent. Hazardous working conditions can also result. Remedial measures, such as scaling large areas and using wire mesh or other artificial support, are very expensive and difficult to implement. There must be a trade-off between the money saved by using larger blasts, and the time spent to maintain pit wall quality. The best approach is to control blasting effects so the inherent strength of pit walls is not destroyed. Presplit and trim blasting are two control blasting techniques used to accomplish this. Both techniques are designed to create a low explosive energy concentration per unit of wall area at the perimeter of the pit. Energy concentration of the main production blast must also be controlled to avoid damaging the final pit wall. A low-energy concentration at the final wall can be obtained by decoupling charges, decking charges, using less powerful explosives, decreasing blasthole diameters, and changing the burden and spacing. Rock properties-fragmentation and the final state of the pit wall-influence the success of a blast. The most important properties to consider are: in situ rock strength, and the nature, frequency, and orientation of structural features. Since these variables cannot be controlled, they must be evaluated by suitable field tests, and then a control blast can be designed using controllable variables, such as spacing, burden, and hole diameter, so the blast will be successful for appropriate rock conditions. One main thrust of this article is to show how control blasts can be successfully designed for larger hole sizes, up to 381-mm-diam. While efforts to produce a good control blast in shovel/truck pits have been ongoing for some time, recent use of a modified presplit technique in dragline operations has helped reduce water problems, improve safety, and increase dragline productivity in a number of surface coal mines. This article presents examples on designing and using presplit and trim blasting for large hole operations in both shovel/truck and dragline mines. Presplit Blasting As introduction to presplit blasting, Fig. 1 illustrates a typical presplit blast layout using 102 mm-diam presplit holes for 381¬mm-diam production holes. For this type of blast, presplit holes would normally be drilled first, ahead of main production holes. The choice can then be made between loading and firing the presplit line or infilling the main blast. In the latter case, the presplit line would be fired instantaneously 100-150 milliseconds before the main blast. As shown in the figure, the presplit line is formed ahead of the main

Jan 2, 1982