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By Russell G. Shomperlen

Grinding mills in excess of 5 meter diameter and 3500 kilowatts per mesh are putting new demands on big gears. This paper returns to the basics of ring gear design beginning with the fundamental power formulas specified by AGMA. Specific emphasis is given to those elements in the rating equations which require closer attention as gear diameters and transmitted power increase. Strict process control in the manufacture of the gear blank is also stressed. Finally, the writer offers a guide to proper installation, lubrication and maintenance.

Jan 1, 1981

By Ramon S. Pizarro

The Masbate Province, Philippines, has been a gold producer for many years. Several gold properties were profitably operating until the outbreak of World War II in 1941. One of these properties was the Masbate Consolidated Mining Company that at the time showed an impressive production rate of more than 4,000 metric tons of ore per day and one of the lowest milling costs on record in the country at that time. Interesting enough, t was the milling equipment left abandoned at Masbate that gave birth to one of the giant copper producers in ASIA. In 1955, Atlas Consolidated Mining and Development Corporation started a new operation at its copper property in Cebu with the seven 2.4m 0 x 3m long (8'0 x 10' long) ball mills left abandoned at the old Masbate Gold Mine. Now, 25 years later, Atlas is processing over 100,000 metric tons of copper ore per day through its three copper concentrators, and, the seven old mills are still grinding, copper ores now.

Jan 1, 1981

By John C. Edwards, Robert C. Bates

Results are given of a mathematical study by Bureau of Mines personnel of overpressurization ventilation effects in underground uranium mines. The mathematics and computer codes make it possible to analyze many facets of transient and steady-state radon diffusion with Darcy flow. Rapid changes in radon flux occur after imposing a pressure differential across the model. Flux into the model mine drops to near zero and then increases to the steady-state level, while the sink flux increases rapidly and then drops slightly to the steady-state level. Magnitudes of mine flux decreases and sink flux increases are dependent upon the distance from the mine to sink, permeability, and amount of overpressure. Current work on more complex models is also described.

Jan 1, 1981

By F. Rosenblum, P. Spira

A laboratory test has been developed for assessing the patential hazard from self-heating of sulphides. The technique consists in measuring the rate of self-heating, from near ambient temperature, of a 1-kg sample in an insulated chamber controlled at standard atmospheric conditions. The rate of oxygen consumption is monitored simultaneously. The paper describes the apparatus and test procedure, and reports data on various ores and mill products. Tailings known to have caused problems underground when used as backfill have shown heating rates of over 3°C/h. A preliminary hazard rating guide for backfill based on the known behaviour of several sulphide materials is presented.

Jan 1, 1981

Develop an environmentally feasible method of mining shallow oil sands without removing overburden. Approach Oil sands are extracted through a single borehole by cutting into the sands around the borehole with a high pressure water jet, and pumping the resulting slurry to the surface. Although oil can be extracted from shallow fields by open-pit mining methods, this type of mining may create significant environmental impacts, including; (1) disruption of the surface, (2) accumulation of waste rock piles, (3) accumulation of tailings, (4) damage to the ground water quality, and (5) surface water pollution. The borehole mining system developed and successfully field tested by the Bureau of Mines offers a method for extracting oil from oil sand with minimal disturbance to environmental quality. (For examples of other applications of borehole mining, see Technology News Nos. 95, 63 56, and 48.) How It Works Prior to the start of mining operations, baseline conditions for ground subsidence and ground water quality are established at the test site so that the degree of subsidence and changes in ground water quality can later be determined. Oil sands mining is conducted using the Bureau of Mines borehole mining system. The borehole mining tool consists of a 12-inch-diameter pipe string and accessories capped with a three passage swivel terminated with a mining section. The mining section has an auger at the bottom. Four slurry inlet holes and an eductor are located above the auger, and a single cutting nozzle is located near the top of the mining section. The tool generates a high velocity water jet that erodes and slurrifies ores. The slurry is drawn into the inlet of the eductor which lifts the slurry to the surface where it is deposited into a slurry discharge sump. To detect evidence of subsidence, surveys of the land surface are made at regular intervals during, and until 30 days after, the termination of the mining operations. Samples of ground water from the borehole and two adjacent monitor holes are taken on alternate days during the mining operations, and are later analyzed to determine if any ground water contamination has occurred.

Jan 1, 1981

By Gunnar Almgren

Hole deviation has great consequences on mining economy, especially for long hole drilling technique. First of all the bore hole datas are concerned but also such factors as development, loss of ore and waste rock dilution are influenced. Hole deviation for sublevel open stoping is illustrated with examples from some mines. The deviations are indicated as mean values but also as standard deviation values, owing to their statistical distribution. The economic influence of the hole deviation is especially studied for steep, small lowgrade orebodies, mined with big hole blasting. Hole deviation can thus influence costs up to US$ 3-4 per ton for operating, development, losses of ore and rock waste dilution. This cost is of the same magnitude as the total operating costs for many sublevel stoping mines. The economic savings in accurate drilling are therefore substantial and are worth all attention.

Jan 1, 1981

By Roshan B. Bhappu, George M. Potter

The chemistry involved in the recovery of gold and silver from natural resources through hydrometallurgical and pyrometallurgical processes is reviewed. Process metallurgy for the cyanidation, activated carbon adsorption, carbon stripping, and electrowinning of gold and silver is discussed. The application of zinc sodium sulfide use for silver recovery from cyanide solutions is covered. Electrooxidation, other oxidation methods and roasting processes for problem ores and concentrates containing organic carbon and impurities such as arsenic, antimony, and tellurides are summarized. The smelting of cathodes, precipitates, and other products, and the various wet and furnace steps in separating dore metal are also reviewed.

Jan 1, 1981

By Norman D. Harte, George W. Holl

The development of the most important tool used in Site Testing is reviewed briefly and the application of the information gained is described. The operation of Site Testing and pregrouting of underground excavations is dealt with in detail. Practical experiences (and pitfalls into which the inexperienced might fall) are recorded. But for the application of Site Testing and grouting of shafts, tunnels and other excavations, vast areas of the present goldfields in South Africa would remain unexploitable.

Jan 1, 1981

By Michael M. Maier, Charles B. Brumleve

The Henderson mine uses the panel caving method to extract molybdenum ore from a deep ore body in an igneous intrusive. Research into caving mechanics and its related problems has yielded fundamental relationships concerning the response of a stressed rock mass to the caving method. The relationships studied through a variety of programs include the following: the evaluation of time and distance related stress fields; the monitoring of the deformation of underground openings and the design implications; ground and mining sequence classifications compared to past support requirements; the response of finite rock masses to zones of high stress concentrations; the evaluation of the subsidence induced by caving; and the microseismic monitoring of cave growth. The cumulative objective of this research is to better understand, predict, and control rock mass response.

Jan 1, 1981

By G. E. McClelland

As part of its mission to assure an adequate domestic supply of metals essential to the Nation's welfare, the U.S. Department of the Interior, Bureau of Mines, investigated a particle agglomeration technique as a means for improving the flow of leaching solution through heaps of clayey or finely crushed, low-grade, gold-silver ores. Bench-and pilot-scale experiments showed that the percolation rate of cyanide leaching solution was markedly enhanced by mixing the ore with a portland cement binder, moistening the mixture, and mechanically agglomerating and aging the feed prior to heap building and leaching. The rate of gold and silver recovery markedly increased because of the increased, uniform percolation of leaching solution through the agglomerated feed. The use of concentrated cyanide solution instead of water during the agglomeration procedure decreased the leaching time required to obtain maximum recovery. Results of bench-and pilot-scale experiments are discussed.

Jan 1, 1981

By Barry James

"This paper discusses past and current energy use patterns in the mining and manufacturing sectors of the industrial minerals industry. Information is presented on the primary energy input types as purchased by the industry, the quality and form of the energy actually required by the various production processes, and the residual form of this energy, either embodied in the product or rejected in gaseous and liquid waste heat streams. With the continuing concerns about the availability of energy combined with the rapidly increasing cost of energy as a production input, the reduction of energy usage through improved efficiency will undoubtedly be an increasing priority for the industry during the 1980s. Reduced energy consumption through investment to improve energy efficiency appears to be advantageous from both the corporate and national perspectives.IntroductionEnergy, like labour and materials, is one of the basic production input s to the various mining, processing and manufacturing production stages in the industrial minerals industry. Energy costs, as a percentage of total production costs, have been well above the average for industry in Canada, accounting for more than 50 cents out of every dollar in the most extreme cases. Consequently, energy has always been a relatively important management concern in the industrial minerals industry.The energy supply and price situation in Canada and internationally has been changing rapidly since 1973, following a long period characterized by declining real energy prices and an abundance and wide choice of energy supplies. It is generally accepted that energy prices will continue to increase rapidly, and that there will be uncertainties about the long-term continuity of supply of some energy types. In essence, a period of relative stability has been replaced by a period of relative uncertainty. In such a changing situation, a reassessment of the role of energy as a production input is appropriate.The following sections outline some of the effects of the changing energy scene on the industrial minerals industry. These effects are dealt with primarily in terms of energy costs, because, in most circumstances, a change in the cost of energy as a production input is the end result of both energy price and energy supply changes."

Jan 1, 1981

By John P. Bernardi, Ralph B. Hall

Past and present phosphate production in the central Florida phosphate district has been predominantly from the unique Bone Valley formation. As these lucrative reserves are depleted, production will shift to the Hawthorn formation in all areas, but particularly so within the southern extension, where vast potential reserves of low grade phosphates exist. Use of existing plants can be extended for some years by mining deeper into the underlying Hawthorn formation, and by mining previously ignored peripheral areas if erratic combinations of the two formations. Hawthorn Production, as compared to the Bone valley, will be characterized by lower grade; lower tonnage yield, lower pebble percent of product, higher mining ratios, higher slimes (- 150 mesh) content, and more complex mining and beneficiation problems. Geological control and more sophisticated operational procedure; will become increasingly more important as the lower quality reserves replace Bone Valley production.

Jan 1, 1981

By Ernest Selig, John Dunnicliff, Delon Hampton

This paper provides a summary of a two-day conference held in New Orleans in March 1980, entitled "Tunnel Instrumentation: Benefits and Implementation." The conference addressed two major questions: First, what are the benefits of using instrumentation during tunnel construction? Second, what contractual relations should be established for conducting instrumentation programs during tunnel construction? Each session consisted of brief formal presentations and extensive panel discussions with participation from attendees. Attendees included owners, designers, contractors, and instrumentation specialists. This paper presents a consensus of views expressed during the conference.

Jan 1, 1981

By S. M. Matthews, D. J. Maconochie, L. G. Alexander, V. H. Tillmann

Case studies of the behaviour of large open stopes at the New Broken Hill Consolidated Mine, Broken Hill and the CSA Mine, Cobar, N.S.W. are considered. Deformations were measured by multipoint rod extensometers in sufficient detail to note the phases of rock behaviour, from intact through to rock-fall. In one case readings were obtained from a zone of stress failure in which two bumps, felt as ground tremors, occurred. In this case noise and stress were also monitored. Deformations were analysed in relation to previously measured in situ rock stress, and conclusions drawn about the degree of conformity to, and the areas of departure from, elastic model behaviour. Attempts to describe the growth of instability utilize reduction factors of the apparent modulus, a magnification factor for the apparent Poisson's ratio, and a correlation between progressive noise count and displacement. Monitoring itself is seen as an aid in determiing the final dimensions of stopes and pillars. One large stope, of intermediate dimension 80 m, was mined essentially without change of plan, whereas a change in mining plan was necessary in other stopes. Aspects of interest to mine planners, and areas where futher research is needed, are indicated.

Jan 1, 1981

Excavation through fill material in underground metalliferous mines is usually avoided where possible, unlike in Civil Engineering application where excavation through soft ground is normally a matter of course. In the mining situation, where excavation is normally for access only, then excavation designs are engineered for mining in hard rock, except where previous lack of forward planning has made such excavation inevitable or a mining strategy sequence necessitates some mining through fill in order to achieve optimum recovery of ore. Historically, this has been the case, because of the lack of knowledge of the support characteristics of the fill material placed underground and the lack of knowledge of the composition of the fill material as placed. The expertise and experience available from mechanised excavation techniques in Civil Engineering application (shield jacking, forepoling, spiling, and soft tunnel boring, to name a few) are available for application to the underground mining situation if the need ever arises. This paper describes the situations in underground mining development where mining through fill material is required, the different types of fill material encountered and the types of support available

Jan 1, 1981

By Raimo Matikainen

Sublevel stoping has been used in Finland since the 1930's and today more than ten under- ground mines have adopted it as their main stoping method (Fig. I). Approximately 80% of the total underground tonnage is mined by sub- level stoping (Table 1) The first mines to introduce the sublevel stoping method in Finland were limestone mines. The development of sublevel stoping was very slow up to the mid 1960's when the switch to trackless mining started and the increase of mechanization lead to higher capacity and productivity. The method itself has been almost the same since the 1950's when long hole drilling was introduced. According to a future prognosis sublevel caving will give way to sub- level stoping; on the other hand, open pit mining will at the same time grow rapidly with some new mineral orebodies, which will he brought into production. Most Finnish orebodies are, however, only marginal ores and they could not support mining with more expensive methods than open pit mining or sublevel stoping. The future of sublevel stoping seems relatively bright. GEOLOGICAL AND ROCK MECHANICAL BACKGROUND The rock mechanical conditions in Finland are in general relatively good and the rocks firm. The compressive strength is normally 40 - 150 Mpa and modulus of elasticity 40 - 200 Gpa. Low values are common in altered and fractured rocks and near the ore contact Rock stress measurements have been carried out already for 20 years for mine design and monitoring purposes. The results of these measurements show that relatively high horizontal stresses seem to be common (Fig. 2).

Jan 1, 1981

By Steve Naruk

Microscopic point-count analyses, although time-consuming, provide detailed mineralogical information which is not obtainable by any other means. Such information can be invaluable for evaluating mill circuit behavior. In comparing two months of operation of a Cu-Pb-Zn sulfide flotation circuit, chemical analyses show that higher copper recoveries and concentrate grades were obtained from lower grade ore, but they provide no explanation of how the improvements were obtained. Quantitative micro-scopic analyses, on the other hand, show the following: 1. There was no change in the identity of the Cu-bearing minerals. 2. Increased copper recovery was achieved primarily by increased recovery of chalcopyrite middling particles. 3. The increased recovery of chalcopyrite middling particles also increased the amount of gangue entrained as locked particles. 4. Increased rejection of free gangue particles, particularly those finer than 25 microns in size, resulted in a net decrease in gangue content.

Jan 1, 1981

By J. Richard Kyle

Lead-zinc ore bodies in the upper part of the Pine Point carbonate barrier complex are localized in paled-solution structures that developed as the result of post-middle Givetian subaerial exposure. Detritus-filled dolines host prismatic ore bodies, and macropores and stratabound permeable zones contain tabular ore bodies. Sulfides are concentrated in these transgressive paleosolution structures that acted as natural bypasses between aquifers and loci for mixing of fluids of different character. Complex ore textures indicate rapid early sulfide precipitation followed by slower sulfide growth. Evidence for both sulfide-carbonate equilibrium and disequilibrium conditions suggest fluctuations in ore fluid composition, perhaps related to periodic supply of reduced sulfur. District and ore body metal distribution pattern may reflect the nature and movement of mineralizing fluids.

Jan 1, 1981

By William L. Barnes, Maurice Leiser

INTRODUCTION The literature on project management considerations and techniques for large rail rapid transit projects in mid-1979 was fairly large and getting larger. UMTA was establishing special organizational elements to deal with the subject as it related to those properties requesting capital assistance, and comprehensive project management plans were being requested from those properties as an early step to build confidence in UMTA that the particular project could be planned, designed, and executed efficiently and effectively, thereby using critically short capital assistance funds to best advantage. In the late 60's and early 70's, the choices of models for heavy rail construction would have been quite limited.., and perhaps much simpler than was the case in 1979. One of the writers was with an agency in 1979 in the process of finalizing its organization for project implementation after several years of study and had the opportunity to review the experiences of others at length, presumably to build on the success and to try to avoid the pitfalls and failures. The properties that were surveyed all reported more or less success stories. One could only conclude that there are many ways to organize for a major transit (construction) project and that there is a great variety of viable management structures, but it seemed that chances for success were best where the executing property had prepared a comprehensive project management plan, then resolved to follow it. Perhaps more important, a commitment or prior approval had been sought from UMTA as to the validity of the project management plan, and the property was allowed to proceed without detailed prior UMTA review or approval of all specific intermediate steps. Most important, however, an early overall UMTA funding commitment had been obtained that permitted some reasonable expectation of the rate of funding, perhaps by commitment

Jan 1, 1981

By Robert W. Miller

There are many problems faced by ventilation engineers in deep underground mining operations, not the least of which is controlling miner exposure to radon gas and its daughter products. Radon gas is commonly found in uranium mining operations, but may also be present in other deep metal mines. For example, tin mines in England, iron ore mines in Sweden, gold mines in South Africa, and molybdenum mines in the U. S. have potential radon exposures. This is because uranium and accompanying radium ore are ubiquitous to the earth's crust albeit at low levels. The fact that the activity represented by one WL can be caused by a relatively low concentration of radon gas increases the difficulty of control. Since the source of the radon gas is usually widespread throughout a mine, local exhaust ventilation is not a viable control schema. The technique used to control exposure is then dilution ventilation and, in fact, huge amounts of air must be moved in order to reduce potential exposures to an acceptable level. An interesting comparison can be made of ventilation rates in different types of mines. It is estimated in modern coal mines, which are generally acknowledged to have high rates of ventilation, that about eleven tons of air are moved for each ton of ore mined. A typical operating uranium mine may have ventilation flows of 14-15 tons per ton of ore mined. This provides an idea of the scope and importance of ventilation in modern mining operations where radon is a hazard. Further pressure is put on ventilation engineers by the steady downward trend in exposure limits set by national and international standard setting agencies. Much of this tendency toward lowered standards is based upon longitudinal mortality studies of miner populations. Another important factor is the limited number of experienced miners available in the labor pool. For optimum production, it is important to have as many experienced miners underground in each shift as possible. However, the average daily exposure in a U. S. mine must be less than .3 WL to permit the miner to work underground for a full year. The ventilation system then must provide enough uncontaminated air to maintain the WL below the .3 TTL level to maximize production efficiency and minimize personnel turnover and the problems associated with it. Ultimately, the goal of the ventilation engineer and health physicist is to protect the working miner from harmful exposures based upon currently acceptable standards. U. S. Federal regulations require that in uranium mines all active work sites must be monitored every two weeks if they measure above .1 WL. Areas that have .3 WL ratios or higher must be monitored on a weekly basis until five consecutive weekly samples show the level has dropped below .3 WL. Also, exposure records must be kept for all individuals exposed to levels exceeding .3 WL. These requirements provide a strong economic incentive to have a ventilation system that minimizes exposure of any personnel. A good ventilation system requires careful planning, operation and backup in order to fulfill its mission of providing adequate clean air. Its proper operation also requires coordination with production personnel so it can be adapted as new areas in the mine open up and old areas are sealed off. The ultimate indicator of ventilation efficiency to control radon daughter exposure is, of course, monitoring working levels. Historically, this has been done using the Kusnetz, Tsivoglou, and Rolle's methods, among others. These methods all require cumbersome equipment and tedious calculations to obtain the measurements that results in WL. More important, however, they require a significant time lag between sampling and counting, typically 40-90 minutes. This time lag is, in fact, what can cause significant economic losses due to unnecessary downtime as well as high WL exposures. In a typical mining situation, a sampling technician using the Kusnetz method takes a sample, moves to the next location and takes another sample and so on. Forty to ninety minutes after the first sample, the technician will stop, run the activity count on the filter and calculate the WL. The technician may be one-half mile away or several levels removed from where the first sample was taken when it is counted. If the WL ratio is high the technician must then backtrack to the sample position. There are then two options. If the sample area is a working stage, it can be shut down or a second sample can be taken. If the first alternative is chosen; i.e., shutdown and correction of the ventilation, then another sample must be taken, followed by a forty minute wait for results. If the ventilation adjustment didn't correct the problem, then the whole process must be repeated with a minimum of forty-five minutes per sample cycle when using the Kusnetz method. It has been estimated from operating uranium mines that the cost per hour for downtime on a production slope is about $1,50O/hour. The time lag between sampling and resultant data can be very costly. If the second alternative is chosen to verify the first reading, the miners may be unnecessarily exposed to high levels while waiting for the result. Clearly, such a sampling system can be markedly improved by eliminating the excessive time lag between sampling and analysis.

Jan 1, 1981