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By J. T. YIi-Penttiia

Dust formation and reaction mechanisms of copper matte particles were investigated in simulated flash oxidation conditions. The granulated and ground matte consisted of 72 wt% Cu, 4 wt% Fe and 21 wt% S. Experiments were conducted by using a laminar-flow furnace with a reaction gas composition ranging from 20 vol% to 75 vol% O2 in nitrogen. The particle feeding was carried out by the single-particle method and samples were quenched after a specified reaction time (approximately 0.05-004 s) in a cold water film. Experimental temperatures were between 800 and 1300°C. Chemical analysis was used to determine oxidation degree, and an optical and scanning electron microscope combined with energy dispersive analysis was used for characterization of the reacted particles.

Jan 1, 1998

By Charles Bruce

Three years ago, at a reception given during the Steel Institute meeting in New York, I heard the remark, "A new formula must be found for foreign investment." This was immediately following the takeover of U. S. Steel's Matilde operation in Bolivia. Recently, at the World Mining Congress in Lima, Peru, the subject of foreign investment was still a lively topic of conversation. The nationalization of the copper mines in Chile and recent trends in the "Third World" have made all but a few extremely wary of investment anywhere other than in the ?stable? countries. Now, the fallibility of this philosophy has been demonstrated in such "stable" countries as Canada, Australia, and Ireland. Yet, nationalization is not new to us. The events in Chile were classified as disastrous to the affected companies, but so were the effects of nationalization to the majors in Bolivia in 1952. Mexicanization of the mining industry in Mexico was considered impractical by many. Now, those who accepted it are finding it to be workable.

Jan 1, 1975

By R. D. Hagni

Pyrometallurgical products from lime roasting of refractory gold ores, pyrite, and arsenopyrite have been examined for their mineralogy, porosity, permeability, retention of contaminant sulfur and arsenic, and potential effectiveness of gold extraction. Cathodoluminescence microscopy, supplemented by X-ray diffraction and scanning electron microscopy-energy dispersive analysis study, has shown that the products phases have characteristic cathodoluminescence properties as follows: anhydrite (CaS04), deep blue; unreacted portlandite (Ca(OH)2), deep purple; oldhamite (CaS), light yellow to rarely pink or orange; a sulfur-rich, arsenic-poor phase, violet; an arsenic-rich, sulfur-poor phase, non-CL. Within the temperature range examined (400-850°C), hematite exhibits greater porosity and permeability at the intermediate roast temperatures of 600-650°C.

Jan 1, 1993

By Lawrence K. Nordell

The selection of belt conveyor drive systems (*) is the subject of ongoing investigation. Frequently this choice has been governed by an incomplete perspective on the major considerations. A more comprehensive and objective approach is desirable. This paper presents additional factors which hopefully will broaden this perspective. BACKGROUND Conventional belt conveyors are powered by electric or electro-mechanical drives(*) mounted at the conveyor's head, tail, intermediate return span, or a combination of these (Fig. 1). A list of the drives reviewed is given in Table I. The following are main elements influencing the selection of the drive system: 1) dynamic characteristics, 2) economic factors such as capital cost, plant spares compatibility, reliability, availability, maintainability, and 3) environmental constraints. The principal dynamic characteristics of concern are the reactions of belt forces generated by the drive system during starting of the conveyor. Another attribute for consideration is the application of multiple driven pulleys used to 1) lower the conveyor's belt tension requirements, 2) increase the expected tonnage throughput, allowing for possible drive component failures, and 3) increase the plant spares flexibility related to the drive assembly.

Jan 1, 1977

By O. H. Tuovinen, T. M. Bhatti

Microbiological leaching has been used as an alternative approach to conventional hydrometallurgical methods of uranium extraction. In the microbiological leaching process, iron-oxidizing bacteria oxidize pyritic phases to ferric iron and sulfuric acid, and uranium is solubilized from the ore due to sulfuric acid attack. If uranium in the ore material is in the reduced, tetravalent form (UIV), a red ox reaction is involved whereby uranium is oxidized to the hexavalent form (UvI) upon dissolution. In acid-leaching systems, the primary oxidant is ferric iron, which is reduced to ferrous iron by its chemical reaction with UIV. The ferrous iron thus formed is reoxidized to ferric iron by iron-oxidizing bacteria such as Thiobacillus ferrooxidans and Leptospirillum ferrooxidans. Nutritional requirements and responses to environmental extremes of acidophilic iron-oxidizing bacteria are appraised. The S-entity in Fe-sulfides is oxidized to sulfate by bacteria similar to Thiobacillus ferrooxidans and Thiobacillus thiooxidans. Pyrite and marcasite oxidation is a sulfuric acid forming reaction. Heap, dump and in situ leach techniques are feasible as bacterial leaching systems.

Jan 1, 2000

By Ramesh Malhotra

Overseas coal trade has increased from 169 to 274 million metric tons (tonnes) between 1970 and 1983, representing a 3.8% average annual rate of growth. Metallurgical coal accounts for nearly 61% of the international coal market while steam coal's share is about 40%. By year-end 1983, the U.S. was the single largest coal exporter at 77 million tonnes, followed by Australia, 56 million tonnes; Poland, 35 million tonnes; South Africa, 28 million tonnes; USSR, 27 million tonnes; and Canada, 17 million tonnes. Coal exports can play a more significant role for one country than another country. For instance, in the U.S., coal exports accounted for only 9% of total coal demand in 1984 while in Australia, exports accounted for 53% of that country's total coal demand. Of all the major coal exporters, only U.S. exports accounted for less than 10% of total consumption in 1984. Japan is the largest importer of coal, 73.8 million tons in 1983. Although the majority of Japan's imports are metallurgical coal, steam coal imports into Japan increased seven-fold between 1978 and 1983, from 2 to 14 million tonnes. Australia is the leading exporter of both metallurgical and steam coal to Japan, accounting for 47% and 55% of Japan's coal imports, respectively. Since 1978, South Africa has developed into a major steam coal exporter to Japan, increasing its market share from 7.9% to 18.8%. The U.S., which was not even a player in Japan's steam coal market in 1978, has made inroads and accounts for 6.7% of Japan's steam coal imports. However, this translates into only one million tonnes. In the European coal market, for the most part, the U.S. is the leading supplier of metallurgical coal and South Africa is the major supplier of steam coal. The U.S. was not even a factor in the European steam coal market in 1978 and has made inroads during the last six years at the expense of Poland and South Africa. From 1978 to 1983, the U.S. has increased its market share in steam coal exports to France from 0.2 to 28%; Spain, 0 to 52%; Italy, 0 to 43.5%; Denmark, 0 to 19%; Germany, 0.5 to 16%; and the Netherlands, 0.6 to 45%, respectively. The use of coal (solid fuels) as a fuel to produce electricity has had increasing signficance for most of the countries analyzed between 1978 and 1983. The flip side to this is that oil has lost its share of this market during this time period. Of all the countries selected, only France reported a decline in coal's share of total electricity production. This is a reflection of France's ambitious nuclear program.

Jan 1, 1986

By R. A. Metz, William H. Breeding

The remarks that follow are based substantially on experience covering 45 years, 80% of which has been in placer work, rather than on a review of available literature. Most commercial placers have been deposited by the action of water. The richer and more difficult-to-mine placers are those in the headwater areas where gradients are steepest. The most lucrative placers are generally in intermediate areas where volumes are greater, fewer boulders are present, and gradients are from 3% to 1-1/2%. The higher volume, lower grade placers are in the lower reaches of river systems where gradients are lower. Where gold-bearing rivers have discharged into the sea, wave action can concentrate values on beaches, past and present. Most of the rich, readily accessible placers were mined by our forefathers. Current opportunities exist: (1) in remote areas where infrastructure has been absent in the past, or development has been prohibited by adverse ownership - political or commercial; (2) in deposits that could not be mined by equipment available to our forefathers; (3) in deposits unidentified by our forefathers; (4) where the-price-of-product/cost ratio is substantially better than in earlier years; or (5) a combination of those factors. When I entered the placer business in the late 1930s, and subsequently, a prevailing opinion believed that glacial deposits should be avoided as irregular in mineral content and composition, and unrewarding to explore and develop; yet an operator has been mining a fluvio-glacial deposit profitably for the past 17 years. Rich buried placer channels, often called paleo-channels were worked in the last century, generally by hand methods, and under conditions that would be unacceptable today. Exploration and mining equipment now available make some of these channels attractive targets. Well-known examples are in California and Australia. The formation of a commercial placer requires a source of valuable minerals. Above primary deposits, there may be eluvial deposits formed by the erosion of gangue minerals and the concentration "in situ" of valuable minerals. Down slope from these deposits are the hillside or colluvial deposits, and below them are the alluvial deposits of redeposited material. Most of the great placer fields of the world are the result of several generations of erosion and deposition. Well-known examples are in California and Colombia. Gold is a very resistant and malleable material, and gold placers may extend for 64 or 80 km (40 or 50 miles) along a river system. Platinum is less malleable, but is very resistant to disintegration. Diamonds are extremely hard, and (especially gem diamonds) may be found over great lengths of a river system. Cassiterite is less resistant to disintegration, and tin placers seldom extend over two miles without resupply from an additional source or sources of mineralizaton. Tungsten minerals are generally more friable, and within a few hundred yards of the source disintegrate to the point that they are uneconomical to recover. Rutile, ilmenite and zircon placers generally result from the weathering of massive deposits, and may be encountered over extensive areas; most are fine grained and durable. What does a geologist or mining engineer look for in placer exploration? The old adage to look for a mine near an existing mine is still valid. You need a source of valuable mineral. Then you require conditions for concentration, which means a satisfactory gradient and/or other conditions that will permit heavy minerals to settle. Nicely riffled gravel, often called a shingling of the bars, is conducive to placer formation. Coarser gravel is logically associated with coarser gold. Excessive clay and/or high stream velocities in narrow channels can carry gold far downstream and distribute it uncommercially over a large area. When material is extremely fine, in situ weathering and concentration become more important. Placers frequently occur distant from lode mines, and one must remember that in a larger watershed the exceptional floods that occur once in a hundred or a thousand years can move great quantities of material long distances. The carrying power of water is said to vary with the fifth or sixth power of its velocity. I am not ready to disagree with Waldemar Lindgren and accept that many commercial placers are substantially enriched by the chemical deposition of gold from solutions; however, I have seen crystalline gold in clayey material quite distant from known sources of primary gold that is dif-

Jan 1, 1992

By Michael J. Howes

The nature of business in a free enterprise system is to maximise the return on an investment i.e. to maximise the wealth of the shareholders. This is normally achieved within constraints dictated by many agencies, one of which is the health and welfare of the employees. In the mining industry considerable effort has been expended in improvements in this field, however, the major problem still confronting the practicing ventilation or environmental engineer is the determination of what standards he should design for, which satisfy both the economic and health requirement s. Normally the basic health requirements are satisfied by the minimum standards defined in regulations and codes of practice. These are, however, based on an acceptance of a certain risk or lower production potential. A prime function of the ventilation engineer is therefore to financially justify and implement improvements to both standards and systems. The main purpose of this paper is to examine the basis and techniques of providing this financial justification. Changes in existing methods to achieve the same standards but at lower costs are also covered.

Jan 1, 1980

By J. G. Groppo

A series of performance studies have been conducted using counter-current column flotation to recover a marketable clean coal product from existing preparation plant fine waste streams. The test work was conducted at several preparation plant sites in Kentucky and Virginia and utilized a 0.15m (6 in.) I.D. column that was 5.79m (19 ft.) high. The refuse streams tested included both thickener feed and classifying cyclone overflow with a typical size distribution of 100 mesh X O. The percent solids of the column flotation feed varied from 1.5 to 15% while the ash content ranged from 30-60%. The concentrates produced contained 2-10% ash with BTU recoveries of nearly 90% using 0.12-1.25 kg/ton (0.25-2.5 lbs/ton) fuel oil and 0.12 kg/ton (0.25 lb/ton) MIBC. These results were consistent with or better than results obtained in preliminary laboratory scale column flotation evaluation. Test results from 7 different seams are discussed as well as scale-up relationships utilized to design a full scale installation presently in operation.

Jan 1, 1990

By John J. Mancini

Elastomeric liner products have been used for over two decades in a wide variety of uses including pond liners, hazardous waste barriers, dam face rehabilitation and mining applications. Asphalt blended with polymers is enjoying renewed interest in the engineering community because of its durability, high yield attributes, fewer CQA steps due to seamless installation, and ease of installation and repair. As in other lining technologies, advancements in polymer chemistry and improvements in installation techniques have had dramatic impact on the quality of fluid-applied asphalt (FAA) membranes. Membrane 6, man­ufactured by Deery Oil, Inc. for over 16 years is a hot-applied polymer modified asphalt membrane that can be applied to a variety of substrates in a reinforced configuration or installed directly on an asphalt surface. By blending thermoplastic polymers with asphalt, very favorable stress/strain characteristics can be achieved (Chambers, 1980). Typically the membrane will yield under rel­atively low stress and also exhibit high elongation capability. In service applications where puncture resistance is a concern, fluid applied asphalt materials have the flexibility and com­pressive strength under heavy loading that leach pad designers may find attractive. Additional security is achieved if long term service life is important as in the case of radioactive waste con­tainment. The Department of Energy selected Membrane 6 over seamed liners for a surface barrier that must isolate and contain radioactive waste for 1000 years. Asphalt has a long history of use as a water-stop material and it is believed that this natural construction material combined with SBS polymers will per-form for a millennium. (Freeman, 1994) The process of installing Membrane 6 varies according to job size and the type of surface to be protected. Lining applica­tions on compacted substrates require the membrane to be applied to a scrim cloth (fig 1). Some early pioneers in FAA technology neglected this step and therefore never achieved per­formance levels that satisfied regulatory guidelines. The scrim provides a cohesive surface that promotes membrane flow while achieving high impermeability (2Xl0 '11).

Jan 1, 1996

By A. L. de Vegt

Hydrogen sulfide is an expensive chemical used in nickel, zinc and copper mining and metallurgical operations for selective recovery and concentration of metals from leach water streams, acid plant blow down streams, refinery electrolyte bleeds and precious metal plant bleeds. Hydrogen sulfide is either produced on site by catalytic reduction of elemental sulfur or transported to the site as liquefied hydrogen sulfide (H2S) or as a sodium sulfide solution (NaHS). Biologically produced hydrogen sulfide has economically very interesting application possibilities in the mining and metallurgical industry. Converting sulfates from waste streams into usable hydrogen sulfide by biological reduction results in operational cost savings in existing lime treatment systems. H2S can be produced biologically for less than $500/ton S. For the past ten years Paques has been engaged in the development and installation of treatment systems based on biotechnological processes to remove sulfur compounds from water, air and gaseous streams. Paques has designed and installed a groundwater treatment system for the Budelco zinc refinery in the Netherlands to remove sulfate and heavy metals. The hydrogen sulfide produced is used for recovery of metals, mainly zinc, from the ground water stream. The system has been operating since May 1992 treating a flow of 5000 m3/d. At Kennecott's Bingham Canyon Utah copper mine a pilot plant started operation in November 1995 to develop the process for selective recovery of copper from leach water using biologically produced hydrogen sulfide from sulfate containing groundwater. Other pilot plants are operated at DOE's Pittsburgh Research Center and at Cominco Research focusing on selective recovery of metals.

Jan 1, 1998

By R. L. Grayson, R. C. Althouse, L. A. Layne, M. J. Klishis

Introduction The researchers combine a detailed microanalysis of accident data and roof bolting field observations to develop new and meaningful risk indices for the roof bolter job classification. These indices incorporate information that has not been used in traditional measures of risk. The significant parameters that these new indices incorporate are the total number of accidents occurring within a particular work activity, the amount of time a miner usually spends performing the work activities in which the accidents occurred, and the average severity of the injuries incurred in the accidents. Being specific for a particular job, they aid comparisons among work tasks performed in the bolting cycle. The indices' specificity provide anew basis for reexamination of training and operational development of miners for specific jobs. Microanalysis The roof bolting microanalysis was performed by examining the empirical data and written summaries for all bolting accidents reported in the West Virginia Safety Information System (WVSIS) for the years 1983-1987. The WVSIS database contained more than 15,000 accident narratives. The microanalysis of bolting accidents covered 2083 reported personal injuries (RPIs). This large number of injuries, performed at an industry-wide level, provides a broad picture of the nature of roof bolting injuries. The analysis focuses on the activities associated with the cyclical routines of the roof bolting process. For this analysis, roof bolting tasks were categorized into work routines based on the sequence of work activities performed by roof bolters. Four work routines were identified within the overall bolting cycle. And these routines comprised 70% (1217) of all roof bolting accidents at the face. The routines include: face area preparation; tramming, positioning and setting the automated temporary roof support (ATRS); drilling holes; and installing bolts. The types of injuries most common within a routine and the average severity of those RPIs were also identified. Table 1 gives a breakdown of the percentage of accidents by the bolting routines. Another 19% of the roof bolting RPIs cannot be classified in a particular routine because of the vagueness of the narratives. However, it was clear that these accidents involved drilling holes or installing bolts. These results are consistent with previous research. It identified drilling and bolting, followed by tramming, as the most frequent sources of injuries (Helander, Krohn & Curtain, 1983). [Table 1-Accidents by bolting routines at the face Accidents Bolting routine Percent Number Face area preparation 18.2 222 Tram, position, ATRS 23.3 284 Drilling holes 34.2 416 Installing bolts 24.2 295 Total 1217] Roof bolter operators were involved in 82% (1706) of the 2083 accidents. The RPIs incurred by classified roof bolters occurred largely on working sections (95%) and at the face area on the working section (75%). Approximately 23% of the 1706 roof bolter RPIs occurred during tasks which were not part of the bolting cycle. These included belt utility work, providing supplies and section moves. Miners in other job classifications who were involved in roof bolting accidents accounted for the remaining 18% of the 2083 roof bolting RPIs. Among this group were section utility men and general laborers (41%), section machine operators (38%), and electricians and mechanics (11%). Another 9% were performing on-section supervisory functions. Face area preparation Face area preparation work led to a substantial number of injuries, although it accounted for the lowest number of injuries of the four routines. Several tasks were associated with the injuries incurred in this routine. Table 2 shows the percentage of injuries associated with face area preparation. [Table 2- Accidents occurring during face area preparation 26.6% Setting temporary and supplemental supports 24.8% Scaling roof and ribs 21.2% Installing ventilation devices 15.8% Handling supplies and materials 11.6% Miscellaneous tasks] The injuries commonly involved overexertion and 'struck by' falling coal and rock. Handling materials often led to overexertion injuries. Nearly 66% of the scaling-related injuries and about 33% of the injuries incurred during the setting of temporary supports involved 'strikes' by equipment, materials, top and rib. Tramming, positioning and setting the ATRS During tramming operations, workers were often struck by components of the machine and hit by loose roof and rib. Poor footing also led to slips, sometimes against the moving machinery. Table 3 gives a breakdown of accidents in this routine. Control levers were a significant problem in this routine.

Jan 1, 1993

By Eric Forssberg, Ming-wei Gao

Four parameters of the Drais stirred ball mitt, i.e., bead load, bead size, flow rate and top feed size have been studied regarding their influence on the specific surface area. A factorially designed scheme was employed for conducting 27 mitting tests. The results show that the bead toad is the most important parameter in terms of increasing the specific surface area with less energy consumption. The bead size is also important. The optimum ratio of bead size to top feed size is found to be 14 to 80, which is much wider than in bat2 milling. The variations in flow rate and top feed size appear to influence the process insignificantly.

Jan 1, 1992

By Bill Dershowitz

Consider a deep gold orebody with grade of 4 g/t (0.11 oz/st). At the time of this writing, with gold prices at historic highs, it could be a good, but not great, prospect. But if it is a vein deposit, and those veins are 200 g/t (5.8 oz/st), it would be significantly more attractive if there was a way to produce directly from those veins, leaving the rest of the rock in place. [Figure 1 The red around each of the four injector wells shows areas of high pressure, where the lixiviant fluid is pumped into the rock mass, with orange and yellow indicating dissipated pressure. The black lines indicate some of the fracture pathways through which the lixiviant, with absorbed metal, flows toward the central recovery well where low pressure is indicated in blue. Lixiviant is pumped out through the recovery well. Image courtesy of Golder Associates.] There is a way to do this, in situ leach (ISL) or solution mining. It has been a viable commercial technology for more than five decades.

Jan 1, 2011

By Steve Kral

One of the world’s most successful mining shovel manufacturers believes its name is not very well known in North America, particularly in the United States. And there are plans under way to improve on that. Never mind that it is one of the largest mining shovel manufacturers in the world. Or that the mining division has 800 machines in operation worldwide, 100 of them in the United States. And the company currently holds 30% of market share worldwide and 10% in the United States. Forget about the fact that it has already sold more machines world-wide this year than it did during 1995. None of that is good enough. O&K Mining GmbH Dortmund is the mining division of Germany’s giant O&K Orenstein and Koppel. It is one of three divisions of O&K Orenstein and Koppel AG, a US$1.2-billion company. The other two divisions are construction equipment and escalators. The mining division, though, is one of the company’s most successful.

Jan 1, 1996

By G. Kordosky, A. Feather, M. Virnig, P. Crane

Introduction The post World War II nuclear arms race and interest in nuclear power generation resulted in a significant demand for uranium in the 1950’s. This led to the development of uranium solvent extraction technology and uranium became the first metal to be recovered on a large scale by solvent extraction (SX) (1), interestingly more than 100 years after the extraction of uranium from an acid nitrate solution using diethyl ether was first reported by Peligot.(2) For a variety of reasons discussed elsewhere (3,4) there have been upheavals in the supply demand picture and price for uranium since the 1960’s. (Figure 1) The low price of uranium from the mid 1980’s until 2004 resulted in a depression of new resource exploration and mine development along with uranium mine / mill closures and decommissioning. People with skills in uranium processing could not find suitable employment and many experienced engineers and metallurgists took jobs elsewhere or retired. This resulted in a significant brain drain with respect to uranium processing technology.

Jan 1, 2007

By M. Z. Mubarok, Z. T. Ichlas, A. T. Sugiarto, K. Sukamto

The use of ozone as an oxidant for the direct leaching of zinc sulfide (sphalerite) concentrate in sulfuric acid medium under atmospheric pressure was explored. The influence of acid concentration, feed gas injection rate, particle size distribution, stirring speed, temperature and slurry density on zinc extraction efficiency was examined. The experimental results showed that the leaching efficiency depends on all of these operating parameters except for stirring speed. It was found that essentially complete dissolution of zinc from the concentrate with the present method can be achieved in only about seven hours under the conditions of sulfuric acid concentration of 2 mol/L, particle size smaller than 74 µm, slurry density of 50 g/L, stirring speed of 420 rpm and feed gas injection rate of 1 L/min at ambient temperature. The experimental results suggest that the dissolution reactions produce independent elemental sulfur that is readily floated and can be easily separated rather than forming a layer on the surface of the reacted particle, as usually observed in a system with ferric-sulfate as oxidizing agent. It was therefore determined that the dissolved ozone play a key role in improving the rate of zinc dissolution from the concentrate. Analysis of the leaching kinetics indicate that the leaching rate follows the shrinking particle model, and the overall dissolution rate of zinc is controlled by a surface reaction.

By H. Boyd

Fluid energy grinding, or jet pulverization, refers to the process of entraining solid particles in high velocity gaseous or vaporous streams and reducing their size by impact and attrition against other particles or some sort of target. Since all desired size reduction is not achieved in one pass through the zone where energy is applied, some form of classification is necessary to return oversize particles to the grinding zone for further size reduction. Applications Fluid energy mills have normally been used to produce products where the particle size is basically finer than 325 mesh. Products having particle size distributions with a maximum size of 4 or 5 µm, and with a high percentage finer than I µm, can be produced using fluid energy mills. More recently, some fluid energy grinding units have been applied economically to produce ground products with 1 or 2% coarser than 200 mesh. It is also possible to consider usage of a jet-pulverizing zone, without the usual integrated air classi¬fier, operated either in open circuit or in conjunction with a screen. Feed sizes to be delivered to the fluid energy mill depend on the material involved, the type of jet pulverizer used, the finished product size requirement, and the upstream process. In some units feed sizes up to l in. or larger can be accepted, but practical economics usually require the starting material to be -4 mesh or smaller and some pulverizers are typically fed at 20 mesh top size. It is not unusual to find feed materials at essentially -325 mesh when extremely fine products are needed, since capacities are improved with finer feeds. Conventionally the fluid energy grinding process has been used for relatively soft and nonabrasive materials. This includes very low¬melting-point and temperature-sensitive materials or coatings, since there is no measurable temperature rise during fluid energy milling. More recently, however, jet pulverizers are being used on highly abra¬sive materials such as silica, feldspar, glass, zircon, calcined alumina, coke, boron carbide, and similar materials. Since the primary energy application is by impact of particles against themselves, there is a low rate of wear resulting in a small amount of product contamination and in low maintenance costs. Usually the fluid energy pulverizer produces a more narrow particle size distribution than other mills provide, and this is useful where such a requirement exists. It is possible, within limits, to adjust classifi¬cation and other functions of the overall fluid energy system to vary the particle size distribution curve and thereby approach the broader distribution obtained from other grinding methods. By controlling the temperature, composition, and humidity of the fluid used, it is possible to incorporate other functions with the size-reduction process. These include drying or adding moisture, coating, mixing and blend¬ing, and even oxidation or reduction and other chemical reactions or their inhibition. The finished product is discharged in the fluid stream and is conveyed pneumatically to the desired point, where a suitable dust collector must be provided. Operation and Mechanical Features One of the earliest fluid energy mills is the Micronizer now manu¬factured by the Sturtevant Mill Co. Similar units are also manufac¬tured by Jet Pulverizer Co. and others. As shown in Fig. 81, the micronizer consists of a shallow circular grinding chamber with noz¬zles arranged at the outer periphery to discharge tangentially inward. Fluid (usually compressed air or superheated steam) is provided to the nozzles from a manifold at pressures normally in the range of 100-150 psig and at temperatures up to 800°F or higher, depending on the raw material, required results, and economic factors. The com¬pressible fluid expands to high velocity through the nozzles and gener¬ates a spiral flow pattern in the grinding chamber, and discharges

Jan 1, 1985

By Susan P. Taite, Alain R. Charest, Peter D. Lewis

The El Sauzal deposit is a classic, high sulfidation, epithermal gold system located in the southwestern Chihuahua state of Mexico (Fig. 1). A Francisco Gold geologist discovered the prospect in1995. El Sauzal was encountered as a result of an extensive grassroots exploration program in the rugged, dissected volcanic terrains of the Sierra Madre Occidental, around the historical Batopilas silver camp. At El Sauzal, high grade gold is hosted within a series of more permeable, low-dipping, lithic and fine-grained dacitic rocks of early Tertiaryage. The mineralization appears tobe structurally and lithologically controlled. The alteration characteristics are typical of a high sulfidation system, including kaolinite, illite, alunite, dickite and pyrophyllite.

Jan 1, 2005

By Ingemar Marklund

INTRODUCTION Today at LKAB, the primary mining method used is sublevel caving, independent of ore-body size, shape, dip, and geographical location. The mining of small ore bodies with relatively flat dip involves a lot of develop¬ment work, a large initial investment, big dilution, and large ore losses. In 1975 serious investigations into other possible methods, especially for those small ore bodies which do not lie in direct contact with other ore bodies, began. Different methods were studied and a number of in¬teresting ideas came from Russian mines. An analysis was made of one of the proposed methods, sublevel benching, in which mining is carried out from raises, and which uses the throw from the blast to place the broken ore. However, there were reservations about working in big open rooms at great depth, and doubts about the possibility of getting acceptable drilling results in an acceptable environment from raised platforms. Testing was therefore delayed for a time. However the investigations continued and in the fall of 1976, it was decided to try vein mining in a small ore body (Indian) (see Figs. 1 and 2). MINING RESULTS The results are now available from the mining of Indian and are as follows: Ore plan area 1200 m2 Level height 50 m Ore volume 60 000 m3 Ore tonnage 270 000 t Iron content 62% Dip 0.69 rad (40°) Ore thickness 14 m Ore compressive strength 26 MPa Ore C-factor (fracture density) 0.21 Hanging wall compressive strength 28 MPa Hanging wall C-factor 0.27 About 40 000 m of the holes were drilled by ring drills with an average productivity of 85 m/shift; loading of a round (2 rows, 750 m of hole) was usually accomplished in one shift. The total ore recovery was about 270 000 t with an average iron content of 59%, and three to four shifts were required per round (4700 t) to load. The principal of "throwing" the rock during blast¬ing worked well, and no ore remains on the footwall. ROCK CONTROL Before mining commenced at Indian, it was dis¬covered that the rock strength in its vicinity was rela¬tively poor. Since the ore is flatly dipping, and mining was to take place with open rooms, the strength of the hanging wall was of great importance. A supplementary program of exploration drilling was done during the fall of 1976 in order to assess fracture spacing and rock strength from the drill cores. Results showed that the strength of the hanging wall was low because of the many fractures. A decision was made to try cable bolting of the hanging wall by extending some of the ordinary blast¬holes. Cables (20 mm diam) were grouted in these larger holes and the rest of the hole was then utilized for ordinary loading (see Fig. 2). About 100 bolts altogether were installed and, to date, only small falls have occurred and the hanging wall is holding. Its exposed area is 2100 m2 DEVELOPMENT WORK On the 350-m level, four crosscuts (two per mining room), one longitudinal drift, and two elongated drifts (one per raise) for the raise climber were developed as shown in Fig. 1. Of the 230 m of drift, 60 m were in ore and the rest in waste. For mining room A alone, a total of 115 m of drift were needed with 30 m in ore. The drift dimensions were about 4.5 X 5.5 m. In the drifting on the 300-m level, two connecting drifts were driven from one existing transport drift forward to the respective raises. The total development was 93 m, all in waste (Fig. 1). The two raises having a length of 80 m each were driven at an angle of 0.68 rad (39°) from the hori¬zontal. They were driven along the footwall contact with a separation of 30 m and had an area of 10 m2. LONG-HOLE DRILLING The drilling of long holes with the ring drilling unit began in the fall of 1977 in raise A. The equipment consisted of an Alimak raise climber, equipped with a lower platform from which the charging was done.

Jan 1, 1982