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By Tim Neil, O&apos

Copper price rebounds A worldwide scramble for diminishing supplies of copper has pushed the red metal's price to more than $2.20/kg ($1 per lb), the highest level since 1980. The first nine months of 1987, world copper production was down 1% from a year ago, while consumption rose 7%. Demand has been fueled by economic strength in several countries, particularly Japan, South Korea, and Taiwan. The price has also been helped by interest rate declines that have prompted manufacturers to build up inventories. In addition, China is expected to be a major buyer in 1988 as it stocks up for its five-year industrialization plan. There appears to be no sign of any significant increase in copper supplies until later 1988 or early 1989. This means supplies probably will remain tight for some time even if world economic activity slows. Strikes and other production problems slowed world output of copper in 1987. As a result, copper supplies are the lowest they have been in 14 years. World inventories currently total 625 kt (690,000 st). That is not much above the 585 kt (645,000 st) level reached in 1974 when the price of copper rose to a record $3.84/kg ($1.74 per lb). This makes all the more interesting what Chile, the world's largest copper producer, has been doing lately. Chile has been the world's largest copper producer since 1982, when it overtook the United States. Table 1 shows Chile's copper grades, reserves, and production at its mines. Chile contains roughly one-fifth of the world's copper reserves. At present production rates, these reserves would last for more than 100 years. In 1986, Chile produced 1.4 Mt (1.5 million st) of fine copper, up 3%. Chile accounted for 17% of the world's copper output.

Jan 1, 1988

By J. E. Gosline

In any theory of a hydrodynamic nature dealing with the vertical flow of gas-liquid mixtures in pipes, the two factors that present the greatest difficulty are the relative motion between the phases and the energy loss resulting from turbulence work. The former has generally been referred to in the literature as "slippage" and the latter as "friction." A third factor of importance is the nature of admixture between liquid and gas for both the relative motion between the phases as well as the turbulence work are functions of the type of mixture. Although much work of a general nature has been performed on the vertical flow problem, there have been very few attempts made to study experimentally the factors involved in the phenomenon individually. Of these, perhaps the most notable is the work of Moore and Wilde1 on slippage. Some writers on the subject have assumed that over a wide range of operating conditions the controlling type of mixture is one in which gas bubbles of uniform size are distributed throughout the liquid, and that the data obtained from experiments on the motion of single bubbles through various liquids may be used in evaluating the slippage. Ver-sluys2 has further assumed another controlling type of mixture in which the liquid occurs as a dispersed phase within the gas and applies data on the fall of drops of water in air to calculating the relative velocity. It was for the purpose of studying the applicability of data on single bubble motion to the slippage factor and to further observe the nature of mixtures that a research program was conducted at the University of California during 1932 and 1933. These experiments formed a basis for

Jan 1, 1936

By J. E. Gosline

In any theory of a hydrodynamic nature dealing with the vertical flow of gas-liquid mixtures in pipes, the two factors that present the greatest difficulty are the relative motion between the phases and the energy loss resulting from turbulence work. The former has generally been referred to in the literature as "slippage" and the latter as "friction." A third factor of importance is the nature of admixture between liquid and gas for both the relative motion between the phases as well as the turbulence work are functions of the type of mixture. Although much work of a general nature has been performed on the vertical flow problem, there have been very few attempts made to study experimentally the factors involved in the phenomenon individually. Of these, perhaps the most notable is the work of Moore and Wilde1 on slippage. Some writers on the subject have assumed that over a wide range of operating conditions the controlling type of mixture is one in which gas bubbles of uniform size are distributed throughout the liquid, and that the data obtained from experiments on the motion of single bubbles through various liquids may be used in evaluating the slippage. Ver-sluys2 has further assumed another controlling type of mixture in which the liquid occurs as a dispersed phase within the gas and applies data on the fall of drops of water in air to calculating the relative velocity. It was for the purpose of studying the applicability of data on single bubble motion to the slippage factor and to further observe the nature of mixtures that a research program was conducted at the University of California during 1932 and 1933. These experiments formed a basis for

Jan 1, 1936

By Samuel Wyer

THE art of natural-gas compressing is now over 25 years old, and has grown at practically the same rate as the increase in domestic natural-gas consumers. There are now over 200 natural-gas compressing stations in North America, aggregating more than 320,000 hp. of compressor capacity and representing a property value of more than $22,000,000, and compressing more than 85 per cent. of all the gas used. The age and magnitude of the art make it evident that the use of gas compressors is a recognized integral part and universal custom of the natural-gas business. The public also is not without its rights and vital interests in this problem. Approximately 1,700,000 domestic natural-gas consumers in North America are dependent upon gas compressors for their natural-gas service. That is, if the use of compressors were to be prohibited, the majority of these consumers would be unable to secure adequate natural-gas service. Each consumer represents between four and five persons, and it therefore follows that the comfort and well-being-as far as natural-gas service is concerned-of at least 8,000,000 persons would be affected if the use of natural-gas compressors should be prohibited. These 1,700,000 consumers have invested, in services, house piping, fixtures, and appliances, an average of about $90 each, or an aggregate of $153,000,000, which is much more than the companies' investment in gas compressors. Furthermore, in all cases where the rate paid by the consumers is fair to the gas company-considering the value of the service rendered by the gas company-the consumers are entitled to continued service and protection of the investment they have made on the faith that the gas service would be continued in the future.

Jan 2, 1916

By H. R. Manouchehri, A. Weh

"High voltage electric pulse comminution (SELFRAG) relies on inducing electrical breakdown that occurs when the resistivity of a dielectric is insufficient to completely block all transfer of electricity while the conductivity is not high enough to fully accommodate the flow of electricity without considerable changes to the crystal lattice. The technology is based on introducing the electrical discharges into solids immersed in water and placed between two electrodes. The technology has received significant attention since investigations have revealed the potential for fully fragmenting particles or pre-weakening rocks in order to reduce the energy consumption in comminution practices, improve liberation and potentially increase the recovery in mineral processing. Mineralogical analysis of samples showed possible liberation at coarser sizes by high voltage breakage in comparison with conventional mechanical comminution.This study explored the potential of implementing SELFRAG to treat the Sandvik’s Mittersill mine tungsten ore (scheelite) for recovery improvement. Accordingly, standard SELFRAG tests were conducted on four samples from the tungsten process plant to evaluate the energy-breakage relation and potential to liberate the value minerals at coarser size fraction for recovery improvement. The size reduction of the ore after processing with SELFRAG revealed that size deportment occurs at 0.4, 0.2, and 0.1 mm, which indicated potential liberation of gangue and target minerals at respective sizes. In comparison with conventional milling, the difference is the deportment of values into 0.4 mm sizes, which was, to some extent, confirmed by SEM liberation analysis. However, more tests are needed to proof the concept. The study has indicated that, despite consuming more energy in comparison with conventional mechanical comminution, there are promising prospects in liberating the values at coarser fraction and improve recovery by making use of SELFRAG. The energy consumption can be reduced through customized generator and equipment setup, which must be further investigated."

Jan 1, 2016

By Pierre Antoine Dhaiti, Gilbert Sergerie, Michel Dagbert, Jean-Maurice Brodeur

"Quality of iron ore is generally determined by several factors which can be correlated or independent. Among these f actors, we generally find per cent iron, silica, magnetite or hardness. Forecasting the short- and long-term fluctuations of these f actors in the future pit is an important feature of mine planning in these deposits. Computer simulation provides a realistic answer to these questions. It allows to generate values for the most important quality features of each shovel load with the proper correlations between them as well as between one load and the next being fully reproduced. Then different sampling schemes f or grade control or blending facilities can be examined on the simulated data. The mechanics of the simulation procedure and its potential applications are described in an example.IntroductionThe purpose of this paper is to illustrate the use of in-situ grade simulation in the planning o f the grade control in an iron ore min e. More precisely, given a production unit of given size, we need to know how many samples have to be collected in that unit so that we ca n recognize its grade with a reasonable precision. Conclusions that we would like to arrive at are of the type: with 4 blasthole samples we ca n get the iron grade of a 10,000 ton bloc k with less than a ± 2% Fe error 95% of the time.The approach t hat we take to solve this problem is simulation. We simulate the grade of samples and the grade of the block that contains these samples. We repeat the simulation for hundred s o f blocks. By comparing the sample grade with the block grade in each case, we determine the actual number o f blocks where the difference sample-bloc k is less than a given limit."

Jan 1, 1983

By J. Nowak

Polish power industry is based on brown and hard coal combustion and transferring the heat into electric power. Over 99% of lignite is used as a fuel in power plants. Lignite is mined on in open pits. Therefore very deep (for example 300 m in Belchatow open pit) excavations having great volume are formed. Thus such type of mining has the impact on land degradation. In order to reduce this effect, at the earliest possible moment the cap rocks are dumped in the pit but at the safe distance from the mining face. Due to the big amounts of lignite delivered to the power plant, there are shortages of solids for filling the pit after completing the mining operations. Geotechnical properties of dumped combustion by-products depend on the lignite type, combustion technology and ashes haulage methods. The paper presents lignite combustion methods used in Poland and their impact on properties of ashes and methods of their transport determining the parameters of dumped wastes. Geotechnical, physical and chemical test results concerning transport and deposition of hydraulically active ashes from the Belchatów power plant are presented. The critical role of geotechnical properties in selecting the deposition technology and post mining management of open pit excavations was showed. The importance of filling the post-mining spaces for restoration the useful function of brown fields and for neutralization of negative effects of mining activity, was underlined.

Jan 1, 2006

By Kirsten Corin, Belinda McFadzean, Cyril O’Connor, Jenny Wiese

As orebodies being mined globally are of ever-decreasing grades, the focus on the concentration process has moved increasingly to the challenges faced in management of the gangue minerals. This is especially so in the flotation of platinum group minerals (PGMs), where head grades are typically in the region of 4–7 g/t. Hence, in PGM operations, a typical recovery of about 80 percent requires depressing about 85 percent of the gangue minerals. In a flotation cell the main control parameters are reagent type and dosage, froth depth and air addition rate. This paper reviews research projects carried out over an extended period aimed at optimizing the depression of gangue minerals: both those that are naturally floatable and those, such as chromite, that are entrained.  This has involved comparing the two major depressants used in the industry — guar and carboxymethylcellulose (CMC) — in the depression of the relevant gangue minerals, as well as investigating their consequent effect on the behavior of the froth phase. Because most PGM flotation plants treat the ore as if all the PGMs are associated with sulfides, copper sulfate is widely used as an activator, but this can also lead to inadvertent activation of gangue minerals. Finally, chromite levels must be kept below 3–5 percent to avoid any negative effects on smelter performance, and thus it is important to reduce chromite recovery which mainly occurs due to entrainment through, for example, froth depth control.   

By Wilbur Nelson

THE year 1923 has been a dull period in oil development in Kentucky and Tennessee, with the exception of the Kettle Creek area along the Tennessee-Kentucky line, in Clinton and Cumberland Counties, Kentucky, and Clay County, Tennessee. The price of oil during the year has been so low that wildcatters have shown little interest in drilling. At the beginning of the year, light oil sold for $2.31 per barrel and medium oil for $2.06 per barrel. Prices increased to the highest point for the year in February, when light oil brought $2.86 per barrel and medium oil, $2.61. In April, prices had declined 35 c. while in May an additional decline of 25 c. went into effect. During June, prices declined 20 c. per barrel; during July, 45 c. per barrel; remained stationary during August and September, and dropped during November and December to the low price of $1.36 per barrel for light and $1.21 for medium oil. At present, an increase of 29 c. has brought the price of light oil up to $1.55, and indications are that the price will continue to increase during the coming months. It can readily be seen that during 1923 the price of oil was such as to prevent any drilling except that necessary to hold leases and to maintain production enough to take care of overhead expenses.

Jan 3, 1924

By A B. Jahed, Ö Aydan, T Ito

In this study, the authors compiled some case history data and present some typical examples of rock slope stability issues in Afghanistan with a special emphasis on Kandahar region and categorise rock slope failures according to the principles of modern rock slope engineering. The slopes considered involve rock-cut, natural rock slopes mainly. The effect of structural geology and discontinuous nature of rock mass on both slope-cuts and natural rock slopes is taken into account and some preliminary assessments on the stable and unstable rock slopes are presented. Furthermore, the seismic effect on their stability is also considered. It is also pointed out that similar issues could be observed in open pit mines, which may be exploited in years to come.

Nov 29, 2022

By M. H. Chetrick

An important factor in the design and operation of a lignite-gasification plant is knowledge of the thermal requirements. These not only vary with different-size plants but also with the operating conditions for a plant of given capacity. Thermal data have been collected since 1945 from a commercial-scale lignite-gasification pilot plant operated by the Bureau of Mines at Grand Forks, N. Dak. The description and operation of this plant and a study of the gas-composition data have already been described in previous reports (1, 2, 3, 4, 5).2) / This investigation was made to review the thermal data taken in tests and to correlate them in terms of the fraction of undecomposed steam and the H2:CO ratio.

Jan 1, 1953

By G. Milojcic

The opencast lignite mines in the Rhinel and Di strict/Germany are equipped with the continuously operating bucket-wheel excavator - conveyor - spreader system. High productivity requires large capital expenditure on main mine equipment as a precondition for efficient mining. In operation, the technology employed leads to a particular cost pattern which is quite different from that of other industries. Labour, auxiliary equipment and mobile vehicles are the most important cost factors which can be controlled by the management. In this respect information processing computer systems are used for systematical and interpretable representation of operational tasks and jobs both in planning and work preparation and in recording the operational data describing the work result and the use of resources. Another aim of the project is to relieve the middle management of major part of its administrative routine work. Job-related work planning and data recording open up various possibilities of analyzing the operational procedures. This analysis shows, for example, what auxiliary machines are particularly suitable for specific tasks, and working groups can be organized so as to take account of the different skills of individual employees for the sake of an optimal work product. Computerized information processing provides the decision makers with topical information related to their respective responsibilities. The mining department of the Hambach opencast mine is the first user of the PERMAD system. Due to the favorable evaluation it has been decided to introduce PERMAD in three other mines in the course of 1992/93.

Jan 1, 1992

By T Mizuta, D Ishiyama

The Vienna Wood hydrothermal is hosted by mid-oceanic ridge type basalt in graben morphology and is located in the Central Manus Basin. The hydrothermal field is about 300 metres in diameter and consist of a sulphide mounds and tall huge death and active chimneys. From the three distinct ore types collected from the Vienna Wood site, the dominant ore minerals recognised are pyrite, wurtzite, sphalerite, chalcopyrite, Fe-oxide and gangue minerals of anhydrite, gypsum, barite and silica. The consistencies in the paragenetic sequence and in the spatial distribution of the sulfides across the three ore types suggest that similar hydrothermal process exist across active and death chimneys. Fluid inclusion studies reveal a homogenisation temperature of 190-260¦C and an average salinity of 5.1-5.4 wt per cent NaCl equiv. The pH of the hydrothermal fluid ranges from 4.2 to 6.1. Bulk chemical analyses indicate that the mineralisation in the Vienna Wood can be classified into the Zn-(Cu-Pb) type of mineralisation. The tenors of the ore and mineral assemblage resemble those of the Mid-oceanic ridge system and ancient ophiolite hosted massive sulphide deposits of Cyprus and Oman.

Jan 1, 2002

By R. Hogg, T. F. Dumm, L. G. Austin, M. Yekeler

"Dedicated to Professor Kurt Leschonski on the occasion of his 60th birthdayAbstractThe kinetics of batch grinding quartz from a feed of 600 by 425 Jim to a product of 80% less than 10 µm have been determined using screening and laser diffractometer sizing for size analysis. The specific rates of breakage decreased by a factor of about three when the material became less than about 100 µm in size, but the primary breakage distribution function also changed to give proponionately more fine material, so that the grinding efficiency expressed as the development of surface area (B.E:T.) per unit of energy input did not decrease. Analysis of the shape of the particles in the 2S x 38 µm size range showed that particles of this size produced by roll crushing or by 8 minutes of grinding of a 42S x 600 µm feed were not different but at long grinding times the particles were rounded. This suggests that the breakage mechanism changes to give more chipping and abrasion and less disintegrative fracture. As the material approached the ultrafine size range it adhered to the mill case and there was no further size reduction. However, a technique for striking the mill case to dislodge the particles was successful in allowing further grinding to 400Jo by weight less than 2 µm.1 IntroductionIt has long been known that it is difficult to accomplish ultrafme grinding in a tumbling ball mill. It has been reported [12, 13] that there is no further size reduction after some hours of grinding cement clinker in laboratory ball mills. Austin and Bagga (1] and Shah and Austin [16] have given an analysis of the kinetics of dry grinding in a laboratory ball mill, introducing the concept of a slowing-down factor K, which is the ratio of the specific rates of breakage at some long time of grinding to the normal specific rates of breakage. They showed that this factor decreased as fine material accumulated in the mill charge. Cottaar and co-workers [8-10] showed that the movement of solid in a dry laboratory ball mill was influenced by the fineness of the solid and by the presence of air, and that evacuation of air from the mill caused particle sedimentation and low rates of breakage.The concept of the slowing-down factor has also been used in wet grinding [17, 18]. At very .high slurry concentrations it appears to be correlated with a steady decrease of mill power caused by the build up of a layer of deposit sticking to the mill case. This reduces the effective mill diameter and, eventually, the balls start to adhere in the layer and centrifuge. A decrease of mill power always indicates less lifting and tumbling of the media and, hence, less breakage. However, slowing-down of breakage rates was also observed at lower slurry concentrations where there was no change in mill power."

Jan 1, 1990

By Amr M. Sallam

In recent years, with the increase of the number of high-rise buildings that are proposed and being built in Central Florida especially in Downtown Orlando, Augered Pressure Grouted Displacement (APGD) piles became a pronounced foundation option. The APGD pile is a bored displacement pile that does not result in drilling spoil. APGD piles are mostly suited for loose to medium granular soils and slightly silty/clayey sandy materials, such as those encountered in Central Florida. The case history presented in this paper involves the APGD piles recommended for a 35-story mixed-use condominium in Downtown Orlando. The subsoil condition was typical for the region; that is loose to medium fine sands underlain by thin layer of plastic sandy clays and plastic clays followed by mixed clayey sands and sandy clays with phosphates and traces of shells transitioned to the weathered limestone. Four APGD test piles were installed and were, statically, tested. Three of the tested piles failed to support the design load. Analysis of load test results showed that the bottom half of the test piles, below the plastic clay deposits, was not effective, which was attributed to structural deficiency resulted from possible pile necking due, in our opinion, to the increase of the porewater pressure within the plastic clays. Three additional test piles were constructed with modified reinforcement cage; that is to extend beyond the plastic clays with tighter stirrups. The additional test piles were, statnamically, tested and were able to sustain the design loads. The construction of the APGD piles through subsoil conditions similar to those of Downtown Orlando, where plastic clayey deposits are encountered, should account for the high porewater pressures that mostly will be developed within the plastic clay, hence, may cause pile necking. Extending the cage to penetrate through and pass the plastic clays reduced the necking potential and improved the test piles performance.

Jan 1, 2006

By L. O. Howard

MUCH work has been done recently on the lead-antimony system1 in connection with lead-rich alloys of commercial importance containing less than 20 per cent. antimony. Dean, Zickrick and Nix have called attention to the perpetuation in the literature of several erroneous data, such as a eutectic temperature of 228° C. at 12.5 per cent. antimony, instead of 247° at 13 per cent. The series is given as a classic example of a straight eutectiferous series in such late books as those of Heyn and Grossmann,2 Jeffries and Archer,3 Guillet and Portevin4 Pulsifer,5 Rosenhain,6 and Liddell,7 as well as the somewhat earlier works of Hofman (1918), Wang (1909) and Williams (1920). The work of Dean has definitely showed a solid solution of antimony in lead at about 2.5 per cent. antimony. While lecturing to a class in general metallurgy at the University of Idaho, the writer had the students prepare a series of lead-antimony alloys for the purpose of illustrating methods of thermal analysis. Proper proportions of powdered antimony (sold as C.P.) and test lead were mixed and melted in graphite crucibles under a reducing cover. Cooling

Jan 1, 1928

By L. O. Howard

Much work has been done recently on the lead-antimony system1 in connection with lead-rich alloys of commercial importance containing less than 20 per cent. antimony. Dean, Zickrick and Nix have called attention to the perpetuation in the literature of several erroneous data, such as a eutectic temperature of 228° C. at 12.5 per cent. antimony, instead of 247° at 13 per cent. The series is given as a classic example of a straight eutectiferous series in such late books as those of Heyn and Grossmann, 2 Jeffries and Archer,3 Guillet and Portevin,4 Pulsifer,5 Rosenhain,6 and Liddell,' as well as the somewhat earlier works of Hof-man (1918), Wang (1909), and Williams (1920). The work of Dean has definitely shown a solid solution of antimony in lead at about 2.5 per cent. antimony. While lecturing to a class in general metallurgy at the University of Idaho, the writer had the students prepare a series of lead-antimony alloys for the purpose of illustrating methods of thermal analysis. Proper proportions of powdered antimony (sold as C.P.) and test lead were mixed and melted in graphite crucibles under a reducing cover. Cooling

By J. G. Craig

This paper initially was to be on the design, construction and operation of the Lucky Friday mill. However, this subject has been pretty well covered in some of the mining publications in the past few years - specifically Mining Congress Journal for April, 1961 and Mining World for March, 1962- : In order to avoid too much repetition, we decided to present in this paper a comparison of the Lucky Friday mill constructed in 1959, and the Mayflower mill built by us in 1962. During the past three years Hecla Mining Company has constructed two relatively small mills: The Lucky Friday mill at Mullan, Idaho 750 T/day; and the Mayflower mill at Heber, Utah - 450 T/day. These mills were engineered, designed and constructed by Hecla, using principally our own personnel, with very little outside (contractor) labor employed. This paper presents a comparison of some of the interesting features of these mills, and where differences occur, it discusses the reasons for the various differences. It compares such points as the testing program, physical layout of the plant, construction costs, the grinding section, the flotation section, tailings disposal, metallurgy, recovery, grades, operating costs, etc. To help illustrate, and to compare these features, we will project pairs of slides showing each of the two mills. First a little background and history on the properties ---

Jan 1, 1964

By A. C. Noble, D. E. Ranta, J. W. Mustard

Skarn deposits present many unique problems in estimating ore reserves, including difficulties arising from the complex character of the mineralization and intricate ore body shapes resulting from structural complexity. Mineralization within skarns occurs in metasomatically altered carbonate horizons produced by magmatic hydrothermal fluids emanating from intrusive stocks. Deposit formation is controlled by both the presence of favorable host rocks and by proximity of these host rocks to a mineralizing intrusion. Ore controls must be defined to account for abrupt changes in grade that occur between miner- alized and unmineralized horizons. Gradational grade changes within the mineralized zones must also be defined in terms of skarn type and included barren rock to assist in reserve estimation. Defining mineralized limits can be further complicated by faulted, rotated, or folded strata affecting the morphology of an ore body. The MacTung copper-tungsten exoskarn deposit, located along the Yukon-Northwest Territories border in Northwestern Canada exhibits a number of unique properties that preclude a simple mathematical approach to ore re- serve estimation. This deposit is used to demonstrate methods for modeling geologic controls and geostatistical properties for a skarn deposit.

Jan 1, 1992

By Golubkov OL, Fedotov PB, Shkylar FR

One of the fundamental parameters of the race- way is its extent (a{ o), which has a significant effect on the distribution of gas and material flows in the lower region of the furnace. pp The known relationships for determination of aC o in most cases link the length of the raceway with the parameters of the blast, while the para- meters of the coke layer, in which the raceway is formed, are either ignored or else expressed in terms which are difficult to determine experimen- tally. Furthermore, the coke combustion propert- ies are not always taken into account in current blast furnace operating practice with combined blast. A complex dimensionless parameter, which re- flects the properties of the interacting media, is the ratio of the kinetic energy of the hot, com- bined gas stream to the potential energy of the coke flow, burning at the tuyere in a gas of known composition. Statistical processing of experimental data on the extent of the raceway zone, obtained from literature and our own work on operational furnac- es, has led to a working formula of the following Form:

Jan 1, 1981