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By C. Richard Tinsley

Like most minerals, coal is inherently a demand-limited commodity. The very sedimentary nature of its occurrence implies greater availability potential than demand. But this situation is overridden by economics among fuels, between coals, and within coal blends. Such considerations make coal forecasting a very hazardous profession indeed. THERMAL COAL If one thought that the lead times involved with a mining project were very long, one has obviously not been exposed to the planning process in the electric generation business - a process seriously confounded by shifts in load growth, environmental pressures, capital intensity, security of fuel sourcing, inter-fuel economics, and so on. But as a general rule, the near-term forecasts for thermal coal can reliably be based on a bottom-up, plant-by-plant analysis. Cement plant conversions can also be reasonably estimated next in order of reliability, although they have a much wider spectrum of coal qualities and fuel sources to choose from with a notably higher tolerance for sulfur and ash. Finally, industrial demand can be assembled from the estimates for conversions by pulp/paper plants, chemical plants, etc. The industrial sector is harder to estimate, since it may involve small boilers or dual-fired units. Assessing demand in the Pacific Rim is relatively a straightforward process in the near term because the major importing countries are all located on the Asian continent with either negligible or very minor (yet stable) indigenous coal production, (itself often operated on a subsidized basis). Furthermore, all imports are seaborne. These major importers are Japan, Korea, Taiwan, and Hong Kong with Thailand, Singapore, and Malaysia up-and-coming consumers. The suppliers to this market all have substantial reserves to back up decades of exports to these countries. Australia, the US, Canada, South Africa, China, and the USSR dominate the supply side. The second oil-shock of 1979/1980 has convinced the importers that reliance on oil can be expensive and eminently interruptible. Thus, they are determined to diversify away from oil' to nuclear and coal for generating electricity and for coal for other purposes where possible. This trend is seen to continue even in the face of the oil glut worldwide and oil-price reductions in early 1982. But the importers are also convinced that reliance on one coal source and, in particular, one infrastructure route for the coal chain from mine to consumer can be equally expensive and interruptible. Strikes in the US and Australia; excessive demurrage at certain ports; relegation of coal to a lower priority on multiple-use railroads in the USSR and China; and concern over escalation on high-infrastructure or high-freight coal chains are among the risks worrying the importers. As a consequence, Pacific Rim thermal coal purchases are being allocated among supplier nations, between ports, and within each country. An example of Japan's shift away from Australia and toward the US and Canada is shown in the estimates in Table 1. But the confidence of the import estimates deteriorates sharply beyond the plant conversion timetables and construction schedules in the near term. If part of the second generation of coal-fired power plants can handle lower-energy coals, the field of suppliers could widen to accept sizeable tonnages from Alaska, Wyoming, Alberta, or New Zealand resources. These supply sources generally have some infrastructure or freight advantage to compensate for their lower quality and to compete on a delivered energy-unit basis. These also offer diversification in sourcing. And the possibility of coal liquefaction in Japan further widens the sourcing network. A great number of Pacific Rim coal forecasts have been generated, especially for Japanese thermal-coal imports which are expected to grow strongly in the 1980's. Since the Japanese themselves have not yet settled their energy policy, the exact numbers are hard to call. Nevertheless, at 50 million tonnes of imports in 1990, Japan would consume 50-60% of the total Asian thermal coal imports as shown on Tables 2 and 6. The next most important consumers are the "island" nations of Korea, Taiwan, and Hong Kong (see Tables 3-5). All three are embarking on power plant developments usually with captive unloading facilities, capable of accepting more than 100,000-dwt vessels. Korea, with no-indigenous bituminous coal, is not especially enamoured with US coals, which are deemed too heavily loaded by freight and infrastructure costs -- up to 70% of the delivered price. Thermal coal contracts are presently split to Australia (70%) and to Canada (30%). Korea Electric Power Co. is already considering second-generation boilers capable of burning lower-quality coals than the present standard. Korea does burn domestic anthracite.

Jan 1, 1982

All companies and individuals who embark on a business venture incur costs. For the most part these costs can be conveniently categorised, firstly into capital and operating costs and then into the various elements of these major cost centres. Although the nature of expenditure naturally differs as between industries certain broad cost types are characteristic of all heavy industries. The capital costs of gaining access to an orebody, equipping a new mine, building a treatment plant and perhaps refining and smelting facilities are likened to capital expenditure by a manufacturer on new plant and equipment required for an expansion or new product innovation. Similarly the day to day costs of labour and materials employed in the running of a mining operation can be likened to similar costs in other heavy industry. There is another group of costs that are not so easily categorised. Most industries incur unavoidable costs that do not form part of the direct or capital costs of physical production. These costs include taxes and other payments to government and research expenditure. The mining industry has a particularly heavy burden of non-production costs and it is the objective here to highlight and analyse a few of them. These costs are exploration expenditure, infrastructure expenditure, royalties and income tax. This is not to suggest that these are the only major non-production costs that the mining industry needs to bear. The industry may incur significant costs for environment protection, land tenure maintenance and on liaising with Government on a wide variety of matters. It is for reasons of subject manageability that the list is confined to the four costs mentioned. For the mining industry, exploration and infrastructure expenditure, royalties and income tax have several distinguishing features. Firstly, they are generally proportionately large costs. Exploration and infrastructure expenditure for example, could well exceed mine development costs. Depending on how royalties are levied, they may be very high in relation to profits earned and many in fact exceed profits. Secondly, except for income tax, they are costs that are generally not payable by other industries. Exploration expenditure is in some respects comparable with industrial research and development. However, because mines are wasting assets exploration is necessary for survival, and it is likely to form a larger part of costs per dollar of profits or sales than R&D. It can therefore be said that exploration is of far more significance to the miner than R&D is to the typical industrial company. Furthermore, unlike exploration expenditure, R&D results are transportable and patents permitting, transferable. The infrastructure burden that miners must carry is unique to the industry. Whereas urbanised industry is not usually liable for the capital cost of power, water, transport, townships and community facilities, the miner must finance these projects himself. Neither the payment of royalties, nor of course, income tax are unique to the mining industry. Royalties are more properly described as a unique charge on some primary industries but however described, they are a charge that manufacturing and service industries do not have to pay. Whilst we all pay income tax, the mining industry's uniqueness derives from the special problem that the industry has in arriving at annual net income in view of the capital intensity of the industry, partly due to expenditure on infrastructure, the indefinite life of capital equipment and its unpredictable residual value, funds expended on exploration and long lead times associated with a mine development. The third distinguishing feature of these costs is that, with the possible exception of exploration expenditure, they are heavily influenced by the policies and actions of Government. As such the industry has a responsibility to promote the industry by emphasising its potential to improve national product including exports and by requesting that the Government provide for the industry in its long term planning and seek its co-operation to prevent the industry incurring inequitable penalties relative to other industries.Finally, exploration and infrastructure expenditure are significant because they are front-end charges on a project. When due allowance is given the time value of money and the deferment of income from a mining development to which they relate, exploration and infrastructure expenditure assume greater importance than the raw money expenditure suggests. For all these reasons, exploration and infrastructure expenditure, mining royalties and mining income tax are worthy of detailed study. This is the purpose of this paper. The objective is to examine the determinants that give rise to them and provide some thoughts on their impact on on-going and potential mining developments and, where relevant, to make some suggestions as to how they may be minimised.

Jan 1, 1977

By R. H. Busch, S. M. Loscutoff, F. T. Cross, R. E. Filipy, P. J. Mihalko, R. F. Palmer

INTRODUCTION During the last decade, several studies in France (e.g., Perraud et al. , 1970; Chameaud et al., 1974, 1979 and 1980) and the United States (e.g., Stuart et al., 1978; Cross et al., 1978, 1980 and 1981) have demonstrated the systematic production of emphysema, fibrosis and tumors in the lungs of animals exposed to radon daughters alone or to mixtures of uranium-mine air contaminants. The studies in beagle dogs have been particularly interesting because of the uncertain etiology of the disease and the (apparently) diverse results of the studies at the University of Rochester and the Pacific Northwest Laboratory (PNL). In the Rochester studies, reported by Morken (1973), beagle dogs were exposed to "normal" room air dust loads and radon daughters from 200 to 10,000 WLM*, delivered in 1 to 50 days (rate of delivery, about 200 WLM per day of exposure). Histological examination of tissues was conducted at 1, 2 and 3 years after exposure for all exposure levels. No cancers were noted in these dogs that received estimated alveolar doses of 34 to 1700 rad (0.34 to 17 Gy). Pathologic changes were found only in the alveolar and bronchiolar regions of the lung. These changes were small, subtle, variable, and widely separated, involving only a very small fraction of lung tissue. Lesions appeared as focal thickening of alveolar septa, with some metaplasia of alveolar cells and some hyperplasia of bronchial epithelium. In the PNL experiments reported by Cross et al. (1978), beagle dogs were exposed in lifespan studies to mixtures of radon daughters (rate of delivery, about 14 WLM per day of exposure), uranium ore dust and cigarette smoke. One group of dogs was exposed to cigarette smoke alone. Except in control and smoke-only groups, the dogs died within 4' years of the first radon daughter exposure, or were killed when death appeared imminent because of pulmonary insufficiency (characterized by rapid, shallow breathing). Control and smoke-only animals were killed at periods corresponding to highmortality periods in the groups exposed to radon daughters and mixtures of uranium ore dust and cigarette smoke. Emphysema and fibrosis were much more prevalent and severe in the lungs of dogs exposed to the mixtures. These dogs also had adenomatous lesions, which progressed to squamous metaplasia of alveolar epithelium, epidermoid carcinoma and bronchioloalveolar carcinoma. Pathologic changes in the upper airways of these dogs were most prominent in the nasal mucosa, and included a few squamous carcinomas in the nasal cavity. Respiratory tract neoplasia was noted after ~4 years exposure and at cumulative exposures exceeding approximately 12,000 WLM. Apart from differences in associated carrier aerosol (room air dust vs. uranium ore dust) and radon-daughter exposure rate (200 WLM/day, shortduration exposure vs. 14 WLM/day, long-duration exposure), the most obvious difference in the Rochester and PNL studies was the observation time following exposure (3 years maximum vs. >4 years). Although neoplasia may not have been observed in the Rochester animals because of the earlier termination of the experiments, it is surprising that other lesions, such as prominent fibrosis and emphysema, were not reported. A follow-up study (reported here) is currently in progress at PNL to determine the pathogenic role of uranium ore dust alone and, in particular, to clarify the role of the ore dust in the production of the massive pulmonary fibrosis observed in the earlier study. Pulmonary function testing (a recently acquired capability) was included in the follow-up study as an indicator of progressive change in lung tissue. MATERIALS AND METHODS Chronic (4 hr/day, 5 days/week) exposures began when the dogs were about 2 1/2 years old. Two identical exposure chambers provided space for simultaneous, head-only exposure of 24 dogs to ~l5 mg/m3 carnotite uranium ore dust. An aerosol diffusion system was incorporated in each chamber in order to channel fresh aerosol past each dog's head; uranium ore dust was added to the inlet room air with Wright Dust Feed Mechanisms* (WDFM). Uranium ore dust and condensation nuclei concentrations were measured daily; chamber aerosols were monitored occasionally for particle-size distributions as described for previous hamster experiments (Cross et al., 1981). The carnotite ore used in these experiments, from the Mitten mine in Utah, was furnished in 1970 through the Grand Junction, CO Office of the (then) U.S. Atomic Energy Commission (now the Department

Jan 1, 1981

By Lloyd H. Banning

ELECTRIC smelting tests were conducted at the Northwest Electro-development Experiment Station, Albany, Oreg., to develop a feasible process for producing silicomanganese from siliceous manganiferous material. The Jackson County, Oreg., material used was massive and contained rhodonite as the most abundant mineral. Both raw and beneficiated materials were tested; various proportions and types of fluxes and reductants were used to make up the furnace charges. The silicomanganese product was used successfully in manufacturing various steels at industrial plants. The manganese content of the samples ranged from 23 to 28 percent; silica plus alumina content was as high as 55 percent-about four times the maximum specification for Metallurgical-grade ore. Sink-float laboratory tests indicated that the manganiferous material was amenable to beneficiation. Over 2 tons of rhodonite concentrate was prepared for smelting tests in a sink-float operation. Only the minus-3/4-inch, plus-8-mesh fraction was fed to a heavy-medium cone; over 97 percent of the manganese was recovered in the sink product, which contained 32.4 percent manganese and 44 percent silica plus alumina. About 25 percent of the silica was removed in the float product.

Jan 1, 1959

By D Maschemeyer

The cobalt production of the Chemical Metallurgical Division of Sherritt Gordon Mines Limited was supplemented, from May, 1962, to June, 1964, by processing 3,500 tons of a calcined cobalt-nickel concentrate. This paper describes the reduction roast -aqueous sulphuric acid leach process and the operating practices employed in recovering cobalt, nickel and copper from this oxidized by-product of a lead-copper operation. Incorporated in this paper is a description of how this process was integrated with the existing ammonia leach -hydrogen reduction process for the recovery of nickel and with the soluble cobaltic ammine process for the recovery of cobalt. Introduction T HE Chemical Metallurgical Division of Sherritt Gordon Mines Limited at Fort Saskatchewan, Alberta, has recently completed ten years of operation utilizing the ammonia leach process for the recovery of nickel, cobalt and copper from sulphide concentrates, matte and other materials amenable to this process. A continuing development program, along with the expansion of certain facilities, has enabled the refinery to attain a production rate of 31, 000,-000 pounds of nickel per year by hydrogen reduction. This represents an increase of approximately 85 per cent over the original design capacity of the plant.

Jan 1, 1965

By W. G. Schultz

"Sylvite's potash mine, which produces 8,000 tons of ore per day, probably ranks as one of the most highly mechanized and automated mining operations in the world. Employing, on a normal 10-hour production shift, one supervisor and a crew of seven operators, Sylvite is able to mine at an average rate in excess of 600 tons per hour. The ore is produced by two Marietta miners weighing 258 tons, the largest and heaviest pieces of equipment ever taken underground. The ore produced by these two Marietta miners is conveyed to the shaft area, cycled through a 15,000-ton storage system and hoisted to surface completely automatically. The operation of all equipment in the mine can be monitored and controlled from a central control panel located underground. Because Sylvite's mine is a new operation with equipment of a new design, the full potential of the operation has not yet been realized. The ultimate capability is clearly evident from the success Sylvite has had over the past year as the average mining capability has risen from 100 tons per hour to 600 tons per hour."

Jan 1, 1973

By C. S. Todd, R. L. McNearny, S. Dun, P. Clark, P. W. Conrad, X. Liu, J. Q. Wu, H. Rhee, W. J. Elliot

Soil erosion is strongly related to soil hydraulic properties. Understanding how surface coal mining affects these properties is therefore important in developing effective management practices to control erosion during reclamation. To determine the impact of mining activities on soil hydraulic properties, soils from undisturbed areas, areas of roughly graded mine spoil, areas with replaced topsoil before seeding, and revegetated areas at the Rosebud Mine in Eastern Montana were evaluated. Field and laboratory studies were conducted to determine the soil hydraulic properties of the four representative areas, including particle-size distribution, organic matter content and saturated hydraulic conductivity. Using this data, Water Erosion Prediction Project (WEPP) simulations and risk analyses were performed to investigate mining’s effects on changes in soil erosion potential. Significant differences were seen in the hydraulic properties of the soils in the representative areas due to the mining and reclamation practices at the mine. The WEPP simulations show that the potential for soil erosion increases as a result of mining activities disturbing the soil. The simulations also indicate that the erosion potential may return to pre-mining levels over time with effective revegetation practices.

Jan 1, 2015

By S. Philipp

This paper describes several variants of placement and design, from a technical and economic point of view, of shafts, ore bunkers and drifts near shafts of a Russian potash mine. Advantages and disadvantages of the position of different workings are shown as well as an analysis of stratigraphic and geological layers and their use for excavating pit bottoms, haulage drifts and bunker installations. The potash mine consists of two hoisting shafts and one service shaft. The hoisting capacity of one of these skip shafts is planned to be approx. 7 Million tons of crude potash salt per year. Both shafts together will have a hoisting volume of about 14 Million tons. When the mine will increase their highest level of production, it will be one of the biggest potash mines in the world. The skips, the hoist drives and the bunker near the shaft will belong to the largest ones worldwide, too. This paper specifies the features of such a double skip shaft pit in detail and demonstrates problems, which depend on the fact that the pit bottom is in a depth of nearly 1200 m.

Feb 23, 2014

By Yongjun Peng, Yufan Mu

Pyrite depression in the flotation of copper ores is difficult due to the activation of Cu2+ ions on pyrite surface. In this study, the radical chain reaction involving the redox cycling of Cu(I/II) induced by sodium metabisulphite (MBS) was exploited to selectively depress the flotation of copper-activated pyrite at pH 8.0. Electrochemical measurements revealed that the presence of Cu2+ and O2 was essential to the formation of SO•− 5 through sulphite oxidation, the main entity to convert the activating species Cu(I)S on pyrite to Cu(II)-hydroxide/sulphate, causing deactivation of pyrite. The depression role ofMBS on pyrite flotation was closely associated with the type of grinding media which affected the electrochemical conditions of flotation. When forged steel was used as the grinding media, a high dosage of MBS was required to depress pyrite, but a low dosage of MBS could almost completely depressed pyrite when high chrome steel grinding media was used. This is because high chrome steel produced a stronger oxidation condition with a higher amount of Cu2+ and O2 which promoted the formation of SO•− 5

Jan 26, 2021

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982

By Stephen B. Castor

The rare earth elements (REE) which include the 15 lanthanide elements (Z = 57 through 71) and yttrium (Z = 39) are so called because the elements were originally isolated in the late 18th and early 19th centuries as oxides from rare minerals. Most REE are not as uncommon in nature as the name implies. Cerium, the most abundant REE (Table 1), comprises more of the earth's crust than copper or lead. Many REE are more common than tin and molyb¬denum, and all but promethium are more common than silver or mercury (Taylor, 1964). Promethium (Z = 61) is best known as an artificial element, but has been reported in very minute quantities in natural materials. Lanthanide elements with low atomic numbers are generally more abundant in the earth's crust than those with high atomic numbers. In addition, lanthanide elements with even atomic numbers are two to seven times more abundant than adjacent lan¬thanides (Table 1) with odd atomic numbers. The lanthanide elements traditionally have been divided into two groups: the light rare earths (LREE), lanthanum through eu¬ropium (Z = 57 through 63); and the heavy rare earths (HREE), gadolinium through lutetium (Z = 64 through 71). Although yttrium is the lightest REE, it is usually grouped with the HREE to which it is chemically and physically similar. The REE are lithophile elements (elements enriched in the earth's crust) that invariably occur together naturally because all are trivalent (except for Ce+4 and Eu+2 in some environments) and have similar ionic radii. Increase in atomic number in the lanthanide group is accompanied by addition of electrons to an inner level rather than the outer shell. Consequently, there is no change in valence with change in atomic number, and the lanthanide elements all fall into the same cell of the periodic table. The chemical and physical differences that do exist within the REE group are caused by small differences in ionic radius, and generally result in segre¬gation of REE into deposits enriched in either light lanthanides or heavy lanthanides plus yttrium. The relative abundance of individual lanthanide elements has been found useful in the modelling of rock-forming processes. Comparisons are generally made using a logarithmic plot of lanthanide abundances normalized to abundances in chondritic (stony) meteorites. The use of this method eliminates the abundance vari¬ation between lanthanides of odd and even atomic number, and allows determination of the extent of fractionation between the lanthanides because such fractionation is not considered to have taken place during chondrite formation. The method is also useful because chondrites are thought to be compositionally similar to the original earth's mantle. Europium anomalies (positive or negative departures of europium from chondrite-normalized plots) have been found to be particularly effective for petrogenetic modelling. REE were originally produced in minor amounts from small deposits in granite pegmatite, the geologic environment in which they were discovered. During the second half of the 19th century and the first half of the 20th century, REE came mainly from placer deposits. With the exception of the most abundant lanthanide el¬ements (cerium, lanthanum, and neodymium), individual REE were not commercially available until the 1940s. Since 1965, most of the world's REE have come from two hard rock deposits: Mountain Pass, United States, and Bayan Obo, China. GEOGRAPHIC DISTRIBUTION OF REE DEPOSITS More than 70% of the world's REE raw materials come from three countries: China, the United States, and Australia. China emerged as a major producer of REE raw materials during the 1980s, while Australian and United States market share decreased dramatically (Fig. 1). Table 2 gives recent annual production figures along with estimated reserves by country, and Fig. 2 shows loca¬tions of significant REE mining. MINERALS THAT CONTAIN REE Although REE comprise significant amounts of many minerals, almost all production has come from less than ten minerals. Table 3 lists minerals that have yielded REE commercially or have po¬tential for production in the future. Extraction from a potentially economic REE resource is strongly dependant on its REE miner¬alogy. Minerals that are easily broken down, such as bastnasite, are more desirable than those that are difficult to dissociate, such as allanite. In general, producing deposits contain REE-bearing min¬erals that are relatively easy to concentrate because of coarse grain size or other attributes. For more thorough discussions of REE¬bearing minerals see Mariano (1989a) and Cesbron (1989).

Jan 1, 1994

By R. O. Williams

It is shown that Ni3Al precipitates homogeneously from nickel-rich alwminum alloys as plates on the (100) planes. Prior to actual precipitation a process occurs which is believed to be one of increasing short-range order. After precipitating the Ni3Al plates enlarge through competitive growth. Discontinuous precipitation can occur simultaneously with the above processes. Recent ideas of the origin of precipitation strengthening appear adequate to explain the hardness changes. REMARKABLY little appears to be known about the precipitation process in Ni-Al alloys in spite of their technical importance. This investigation originated to supply additional information about precipitation in general, this system in particular. Information on the structures and kinetics have been obtained through the use of hardness, X-rays, microscopy, calorimetry, and resistivity on high-purity alloys. PROCEDURES Six alloys, Table I, were prepared by melting carbonyl nickel and high-purity aluminum in alumina crucibles in vacuum and casting into 1-in. graphite molds. All rods were homogenized at least once at 1300°C for 24 hr prior to swaging and this was repeated on the first three alloys after 75 pct reduction. Alloy 4 could be reduced only 10 pct at 1000°C (probably in two-phase field) prior to fracture but 1/4-in. samples quenched from 1100°C were readily reduced cold. Alloy 5 was reduced 15 pct cold but failed on the next pass while alloy 6 of essentially the same aluminum content failed inter-granularly without apparent flow up to 1000°C. The alloys were heated in hydrogen at the elevated temperatures and formed thin, coherent aluminum oxide coatings which provided excellent oxidation resistance at lower temperatures. However, freshly prepared surfaces showed considerably less resistance at 500"to 700°C in air and apparently resulted in internal oxidation. As a consequence, low-temperature agings were carried out in evacuated tubes. RESULTS The isothermal hardening behavior of these alloys at 500"and 565C is given in Figs. 1 and 2. These results were obtained from samples cold worked 75 pct, recrystallized at 1000°C (1100°C for the 7.8 pct Al) and quenched in water. This recrystallization was used to give smaller grain sizes so as to obtain more uniform hardness values and the points represent an average of five readings. The electrical resistivity was measured on 1/16-in. wires quenched from 1000°C during aging at 495°C to give Fig. 3. The energy release and its rate are given in Fig. 4 for the 6.9 pct Al alloy during aging around 500°C. Inasmuch as this was a single run, its accuracy is not known but certainly the general shape and magnitudes are correct. The method used to obtain these results is described elsewhere.' Data for the aging at 600°, 700°, and 800°C of these alloys cold worked 50 pct are given in Fig. 5. Supplementary information from microscopy and X-ray diffraction have been included to indicate recrystallization, discontinuous precipitation and the appearance of superlattice lines from the Ni3Al. The hardness of these alloys as annealed, aged, cold worked, and cold worked and aged is given vs composition in Fig. 6. Those samples which were isothermally aged, Figs. 1 and 2, were reaged at 532°C and at successively higher temperatures for the indicated times to give the data of Fig. 7. These results as well as certain others, support the idea that the level of hardness reached for temperatures above 600°C are equilibrium values more or less independent of path. This being the case, the breaks in the curves would be the complete solution of the Ni,Al. The electrical resistivity versus temperatures for some of these alloys, both aged and unaged, is given in Fig. 8 along with those data from heating slowly (10 deg per day) to high temperatures. Interesting points include the lowering of the Curie temperature (the change in slope), the lack of any indications of a solubility limit and the large temperature coefficient for the Ni3Al. A slight break for Ni3Al around 100C shows up but this is not a Curie temperature as Ni3Al is not ferromagnetic down to -190°C. Metallographically both the nickel-rich solid solution and the Ni3Al appear very much like pure nickel. Profuse twin boundaries are present both

Jan 1, 1960

By K. T. Aust

It is proposed that the intergranular corrosion of austenitic stainless steels is associated with the presence of continuous grain houndary paths of either second phase, or solute segregate resulting from solute-vacancy interactions. Experimental observations of structural changes and crrosion behavior of different types of austenitic stainless steel provide support for this poposal. On the basis of this model, it is shown that the intergranular -corrosion susceptibility of austenitic stainless steels in nitric-dic hromate solution may be substantially reduced either by suitable heat treatments or by impurity control. AUSTENITIC stainless steels, such as Type 304, generally have excellent corrosion resistant properties when properly solution heat-treated and used at temperatures where carbide precipitation is slow. However, several corrosion environments have been found which produce intergranular corrosion of solu-tion-treated stainless steels, that is, those steels with no detectable carbide precipitation.''2 Of the various corrosion environments, the most widely used test solution has been the boiling nitric-dichromate solution. In these acid solutions, stainless steels have been found to be susceptible to intergranular attack despite the addition of carbide-forming elements such as titanium or columbium, or despite lowering of the carbon content or use of high-temperature solution treatments. Studies of the electrochemical mechanism of corrosion attack have been made by several worke1s3'4 who found that oxidizing ions such as crt6 depolarize the cathodic reactions and consequently raise the open-circuit potential of stainless steel immersed in nitric acids. As a result of this, the anodic reaction is accelerated. The reason for the localization of anodic activity at the grain boundaries, and resulting intergranular corrosion, has not been conclusively determined. Several workers, e.g., Streicher,3 and Coriou et al.,4 have suggested that the strain energy associated with grain boundaries provides the driving force for the accelerated intergranular corrosion. This argument would predict that alloys of high purity would still be susceptible to intergranular attack. However, work by chaudron5 and by ArmijO,6 has shown that high-purity alloys are immune to attack, in disagreement with this argument. An alternative suggestion is that chemical concentration differences exist between grains and grain boundaries, that is, impurity segregation at boundaries, and that these chemical differences provide the driving force for localized attack. It is this impurity segregation which can lead to accelerated dissolution of grain boundaries when the alloy is exposed to a suitable corrodant. This mechanism would predict the immunity of high-purity alloys to inter-granular attack, which is in agreement with experi-mental observations. In the present paper, some recent studies on inter-granular corrosion of austenitic stainless steels which were conducted by coworkers and myself will be re-tibility A simple model will be described in which it is proposed that the intergranular corrosion of aus-tenitic stainless steel is associated with the presence of continuous grain boundary paths of either second phase or solute-segregated regions.* On the basis of this model, it is suggested that the intergranular corrosion rate can be markedly reduced by the formation of a discontinuous second phase at the grain boundaries if the discontinuous second phase incorporates the major part of the segregating solute, drained from the grain boundary region. Results are presented of corrosion tests and electron microscopic studies of different types of austenitic stainless steel after various heat treatments which provide experimental support for this model. Finally, a solute clustering mechanism, based on a solute-vacancy interaction, is shown to be consistent with the results obtained for inter-granular corrosion of solution-treated austenitic stainless steels. EXPERIMENTAL Corrosion tests using weight loss measurements were made on sheet specimens, which were lightly electropolished, washed, and immersed in boiling (115°C) 5 N HN03 containing 4 g crt+6 per liter added as potassium dichromate. Studies in which the inter-granular penetration depth was measured both by electrical resistance and metallographic methods have shown an empirical correlation between the rate of intergranular penetration and the weight loss per unit time for identically treated specimens of stainless steel." As a result, although all the corrosion data reported here are in terms of simple weight loss measurements, these data are considered to reflect primarily the rate of intergranular dissolution. Fig. 1 shows a typical result of intergranular attack of a solution-treated Type 304 stainless steel after 4 hr in a boiling nitric-dichromate solution. The wide grain boundary grooving at the surface, and the attack at incoherent twin boundaries, are evident; very little corrosion attack is seen at the coherent twin boundaries. INTERGRANULAR CORROSION MODEL

Jan 1, 1970

By Yi Luo

Bolting the mine roof is one of the noisiest underground mining operations. The main contribution to the noise is the drilling of the roof bolt holes, especially drilling in hard rock. Previous research shows that the specific energy consumed in drilling bolt holes is inversely proportional to the bite depth (i.e., penetration per drill revolution). Less specific energy at a reasonably high bite depth means less energy is wasted in the drilling process for producing heat, bit wear and noise. Therefore, proper control of the drilling operation to achieve a reasonably higher bite depth can reduce drilling noise from its sources as a proactive noise control approach. In order to explore the feasibility and develop the proactive technology for reducing drilling noise, drilling tests have been conducted in laboratory settings. The preliminary results show that significant reduction in noise dose can be achieved by drilling in medium hard rock with uniaxial compressive strength of about 276 bars (4,000 psi) and at a reasonably high bite depth of about 4.6 mm/rev (0.18 in./rev). This paper presents the findings from laboratory experimental studies.

By J. H. Jackson, P. D. Frost, C. H. Lorig, L. W. Eastwood, A. C. Loonam

It is well known that for equal weights of material, thin sections of the lighter structural alloys are more resistant to buckling under a compressive stress than thin sections of more dense material. Therefore, structural parts having the same resistance to buckling stresses are generally lightest when they are made of magnesium alloys. Consequently, there is considerable interest in the development of strong alloys having densities equal to or lower than that of magnesium and strength-weight ratios equivalent to those of the strongest aluminum alloys. The development of commercial magnesium-base alloys has been very successful, and their importance among materials in the structural field has been amply demonstrated. However, an even wider structural application of magnesium alloys might be effected if improvements could be made in the cold rollability and cold-forming characteristics. Such improvements, along with production of less directionality in properties. could be effected if the hexagonal close-packed lattice of present alloys could be replaced with a cubic lattice. In 1942, a project, having as its objective the improvement of some of the characteristics of commercial magnesium-base alloys, was initiated at Battelle Memorial Institute under the sponsorship of the Mathieson Chemical Corporation. At that time, one of the authors,* then Chief Metallurgist for the Mathieson Chemical Corporation, postulated that the addition of lithium to magnesium in sufficient quantities to change the crystal structure of the resultant alloy from a hexagonal to a body-centered cubic lattice should produce a magnesium-rich alloy which would have the desired improvements in cold-working characteristics with less directionality in properties. To test this view, a series of alloys was made and was found to have many of the attributes that were predicted. The view that magnesium-lithium alloys were of structural interest was also held by others. In 1943 and 1945, Dean and Andersonl,² obtained patents on magnesium-base alloys containing from about 1 to 10 pct lithium, from about 2 to about 10 pct manganese, and the balance substantially all magnesium; and on magnesium-base alloys containing from about 1 pct to about 10 pct lithium, from about 2 to about 10 pct manganese, from about 0.5 to 2 pct silver, and the balance substantially all magnesium. They noted that an alloy containing 83 pct magnesium, 10 pct manganese, 5 pct lithium, and 2 pct silver could be cold rolled by any of the usual methods and that this alloy was considerably harder and stronger than most other magnesium-base alloys heretofore available. In 1945, Hume-Rothery, et al.,3 predicted that magnesium-lithium base alloys should be soft and ductile and that such compositions, to which was added a third element for the purpose of producing a precipitation-hardening type of alloy, should be ductile, strong, and lighter than magnesium itself. Hume-Rothery investigated the binary magnesium-lithium and ternary magnesium-lithium-silver equilibrium relations and criticized the existing equilibrium diagrams. The binary magnesium-lithium equilibrium diagram has also been investigated by Grube, Von Zeppelin, and Bumm, Henry and Cordiano, Sal'dau and Shamrai, Hof-mann,' and Shamrai.8 The work of most investigators agreed with that of Grube, whose constitution diagram is shown in Fig 1. From the standpoint of development of structural alloys, the room-temperature equilibrium relations are of great interest. An examination of Fig 1 reveals that, from 0 to 5.7 pct lithium, the existing phase is alpha, which is a solution of lithium in hexagonal magnesium. The boundary between the alpha, and alpha plus beta regions, occurs at a Mg/Li of 16.5, corresponding to 5.7 pct lithium by weight. Between 5.7 pct and 10.3 pct lithium, there exists a mixture of the alpha phase (lithium dissolved in hexagonal magnesium) and the beta phase (magnesium dissolved in body-centered-cubic lithium). The boundary of the alpha plus beta and beta regions occurs at a Mg/Li of 8.7, corresponding to a lithium composition of 10.3 pct. Much of the work described here was directed toward obtaining an alloy made up of a large proportion, or entirely, of the body-centered-cubic beta structure. In his early work, Loonam found that specimens of lithium-bearing alloys were very malleable. Ingots of alloys prepared at that time were extruded to 3/8-in. diam bars, and the mechanical properties and the effects of heat treatment were then determined. These data are given in Table 1. The outstanding ductility of the alloys, combined with their moderately high strength, evoked considerable interest.

Jan 1, 1950

By C. A. Rowland

Introduction Rod mills are horizontal, cylindrical, rotating, tumbling mills in which steel rods, that are usually about 6 in. shorter than the length of the grinding chamber, are used for grinding media. Fig. 33 shows a cutaway view of a rod mill. Rod mills were first used around 1910 because of their ability to grind in open circuit, consistently produce a uniformly sized prod¬uct, give greater continuous availability, and give lower maintenance costs. They have basically replaced roll crushers and coarse grinding ball mills in comminution flowsheets. The present 20-ft limit on rod length, based on rod manufacturing methods, and the pulp flow char¬acteristics, that influence mill capacity, have established the maximum rod mill sizes' of 15 x 21 Vs ft using 20-ft rods and drawing from 2,200 to 2,300 hp. Rod Mill Uses. The principal uses for rod mills are: preparing feed for ball or pebble mills, grinding gravel or aggregates to the size analysis required for making concrete sand, preparing coarse agitator feed for leaching operations, and for primary grinding circuits, particularly in industrial mineral applications. Wet grinding is the most frequently used grinding process for rod milling. There are a few special cases where dry rod milling is used. Preparing feed for ball or pebble mills is the most common use for rod mills. In ferrous and nonferrous ore concentrators, wet open-circuit grinding is used to prepare ball-mill feed. Figs. 34 thru 37 show the four basic grinding circuits used in these concentrators. In rod mill-ball mill grinding circuits, where the rod-mill discharge contains 25% or more of the grinding circuit product-size material, the rod-mill discharge can be fed to the ball- or pebble-mill classifier as shown in Fig. 34. This is the most frequently used rod mill-ball mill circuit in the nonferrous ore concentrators. In a few concentrators such as Anaconda's Weed concentrator,55 the rod mill discharges to one classification stage which makes the same or a slightly coarser particle-size cut as the classifiers used to close the ball- or pebble¬mill circuits, as shown in Fig. 35. In concentrators where rod mill-ball mill circuits are used to grind low grade magnetic iron ores such as Reserve Mining Co., Silver Bay, Minn.; Erie Mining Co., Hoyt Lakes, Minn.; Eveleth Taconite Co., Forbes, Minn.; US Steel, Mt. Iron, Minn. and Atlantic City, Wyo.; Inland Steel-Jackson County Iron Co., Black River Falls, Wis.; and The Cleveland-Cliffs Iron Co., Adams mine, Kirkland Lake, Ont., the rod mills discharge to rougher magnetic separators (cobbers), the concentrates from which are fed to a ball milling stage as shown in Fig. 36. The same basic circuit is also used with gravity concentra¬tion methods such as spirals (Hanna Mining Co., Groveland, Mich.), tables (St. Joe Minerals, Bonne Terre, Mo.), and jigs. When the rod-mill discharge contains less than 25% of the grind¬ing circuit product size, this indicates an ore that does not readily produce the fines needed to help carry the coarse feed into the ball charge. When this occurs, the fines in the rod-mill discharge are needed in the ball mill or pebble mill so the rod-mill discharge flows direct to the second-stage mills as shown in Fig. 37. Rod mills are used to make feed for open-circuit ball mills such as those used in wet grinding cement raw material (American Cement, Brennan Ave¬nue plant, Detroit, Mich., and Port Huron, Mich., plant, and at the South Dakota State Cement plant, Rapid City, S. Dak.). Grinding gravel or aggregates to the size analysis required for making concrete sand is used in areas where natural sand of the proper particle size gradation required to make concrete is not availa¬ble. Three-quarter inch crushed aggregate is ground to 4 mesh using center peripheral-discharge rod mills (Fig. 38) operating in open circuit. US Corps of Engineers specifications state that the sand product be 93-100% passing 4 mesh and contain no more than 3-7% passing 200 mesh. It may be necessary to classify the rod-mill discharge to remove any excess -200 mesh sand produced in the rod mill. Particu¬larly when processing aggregate with high abrasion indices, center peripheral-discharge rod mills are used instead of hammer mills, fine bowl crushers, and crushing rolls for the production of specification sand. If the -4 mesh size fraction of the rod-mill feed is removed

Jan 1, 1985

By K. E. Stanfield

This report presents the results of analytical studies made on two series of oil-shale samples from the Mahogany ledge of the Green River mine the character of the various grades of oil shale that are being mined and processed into shale oil and refined products at the Bureau of Mines Oil-Shale Demonstration Plant at Rifle, Colo. The analytical work was part of a synthetic liquid fuels program being conducted at the Petroleum and Oil-Shale Experiment Station, Laramie, Wyo. The 16 oil shales ranG6d in color from a gray brown ( or a sample that assayed 10.5 gallons of oil per ton to a very dark brown for a sample that assayed 75.0 gallons of oil per ton. In general, the samples were similar as to the types of organic and inorganic matter that they contained but varied widely in the amounts or these constituents. The principal constituents were calcite, dolomite, illite, and yellow organic material, while the minor constituents were quartz, feldspar and plagioclase feldspar, pyrite, marcasite, analcite, opal, black organic material, and small woody fragments. Some micro-fossils, spores and pollen grains were found also.

Jan 1, 1951

By Jack E. Bailey

From a start in 1957, milling 3,000 TPD of copper ore, the Cyprus Pima Mining Company, which is located 20 miles south of Tucson, has expanded its operations four times, and is currently treating over 55,000 TPD of ore.1 The most recent expansion added a 14,000 TPD semi-autogenous concentrator to the operation. Shortly after startup in January, 1972, the semi -autogenous plant exceeded the designed tonnage rate.2 Ore from the mine is delivered to a 54 inch gyratory crusher and is reduced to a minus 8 inch product. From the crusher the ore is stacked on a surge pile and then reclaimed and fed to the two parallel grinding circuits. A 28 ft. diameter by 12 ft. long semi-autogenous mill and a 16.5 ft. diameter by 19 ft. long ball mill make up each grinding circuit. The semi-autogenous mills are in closed circuit with screens and the ball mill circuits are closed with hydro- cyclones. Rougher flotation follows the grinding operation. Concentrate regrinding and cleaner flotation is performed in Mill I, the older mill. The generalized flowsheet is shown in Figure 1.

Jan 1, 1975

By R. L. Ash, R. R. Rollins, C. J. Konya

An investigation was performed to determine conditions for optimizing the spacing of simultaneously initiated multiple explosive columns. This was done by using models of mortar, dolomite, and Plexiglas with 10-grain mild detonating fuse as the explosive charge. It was desired to simulate blastholes with multiple primers initiated by detonating fuse or when high-velocity explosives are used in low-velocity materials. It was found that optimum spacing between multiple charges was strongly influenced by charge length. At less than optimum charge length, the spacing at which complete shearing was possible between adjacent charges decreased exponentially with a subsequent loss of broken material volume. For charges fired simultaneously, larger burdens and spacings were possible as compared to those necessary for single-crater charges. For each material studied, there was a characteristic optimum charge length and a maximum attainable spacing at any given burden. Proper selection of the spacing distance between charges is fundamental to successful blasting. Its value directly affects the cost of drilling and explosives used per unit of broken material. In addition, the choice of a spacing that is Compatible with a given set of blasting conditions aids in the control of fragmentation sizing, ground vibrations, overbreak, and throw which in turn, influence other production costs. For example, normally loaded blastholes that are spaced too closely invariably promote overbreak and usually give coarse fragmentation. Unless care is taken, airblast and violent flyrock will occur and under certain conditions cutoffs and misfires may result. Too large a spacing, on the other hand, frequently leads to conditions that form bootlegs or toes. The choice of a particular spacicg to use, however, is largely a matter of individual experience and judgment, usually based on trial and error. Very little is known or can be found in the literature with regard to how the spacing between charges is related to field conditions and charge geometry. As a general rule, the firing time sequence of adjacent charges and properties of a material are thought to have the most significant influence on the spacing distance best suited for any given field condition. For example, delayed initiation of adjacent charges usually always requires a closer spacing than when charges are fired at the same time. This should be expected if one considers that the energy normally dissipated and lost in the surrounding ground from charges fired independently would be captured and utilized for breaking material between charges when they are initiated together. Spacing can be extended also when charges are aligned with structural planes of a material, such as jointing, along which shearing is relatively easy. It is customary to relate the spacing (S) between charges to their common burden (B) in the form of a spacing ratio, or SIB. The burden normally is considered as the optimum depth or distance from any single charge perpendicular to the nearest free or open face at which the desired fragmentation and maximum crater yield are obtained. For production blasting, value of the ratio is generally considered to vary from 1 to 2, depending on conditions.1-6 When adjacent charges are fired independent of one another, the value varies from 1 to about 1.4, the closer amount being employed to square corners or produce craters having the ideal 90" apex angle. The larger ratio is the geometric balance value for craters having an apex angle of 135". The basic ideal crater forms in the plane of the charge diameter for charges fired independently are shown in Fig. 1. In the event charges are fired simultaneously, geometric balance in the plane of their charge diameters suggests that a spacing ratio near 2 would be appropriate, as illustrated by Fig. 2. In practice, however, some compromise ratio value must be selected to conform with the specific ground conditions. An example would be where the jointing planes tend to produce 60° or 120° crater angles, the appropriate geometrically balanced charge arrangement being given by Fig. 3. In this condition, the spacing ratio is 1.15, not 1 or 1.4 as suggested for the 90° cratering of independently fired adjacent charges. In view of the foregoing, it would seem logical to assume that whenever charges all having the same burden are fired at the same time, spacing distances always can be greater than those permitted by charges fired independently. In practice this is not the case, however.

Jan 1, 1970

By L. Pan

As an air bubble approaches a mineral surface, a thin liquid film (TLF) is formed in between. The excess pressure (p) in the film determines the kinetics of drainage, which is of critical importance in flotation. We have developed a new method of monitoring the kinetics for both stable and unstable wetting films and analyzed the results using the Reynolds lubrication theory. It has been found that initially a TLF is thinned by the Laplace pressure (pcur) due to changes in bubble curvature and subsequently by the disjoining pressure (?) created by the surface forces present in a wetting film. In the present work, we have derived an analytical expression for determining ? from the temporal and spatial profiles of TLFs on mineral surfaces. The results show that ? > 0 in a film formed on a hydrophilic surface and ? < 0 in a film formed on a surface hydrophobized by collector. Thus, the role of collector in flotation is to create a hydrophobic force and hence a negative disjoining present in wetting films. It is also found that smaller bubbles can create higher Laplace pressure, which helps the TLF drain faster.

Sep 1, 2012