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By Phillips H. R

The action of free rolling disc cutters is well understood, relationships between disc geometry and performance having been established both in the laboratory and under field conditions. However, under severe operational conditions a malfunction of the disc bearing assembly is likely to produce a definite change of cutting action. Collapsed or siezed bearings can even prevent the disc from rolling and thereby create what is equivalent to a drag pick. This paper describes a programme of laboratory experiments using deliberately stalled discs. Their performance under such conditions is contrasted with that for a free rolling disc and also with a drag pick having a ridged front developing into a negative radiused back clearance angle. In conclusion an assessment is made of the influence stalled discs can have on the torque and thrust requirements for a tunnel boring machine.

Jan 1, 1981

By J. Eric Belford

The Kidd Creek orebody is a massive base metal sulphide deposit with surface dimensions of 168 m by 670 m. The steeply dipping ore- body has been evaluated to a depth of 1 524 m. To date, the mining of the orebody has been divided into three stages. The first stage of mining was an open pit operation which supplied ore from 1966 until 1977 when a depth of 219 m was reached at the widest or north end of the orebody. The second stage of mining, called No.1 Mine, was the development of an underground mine to replace the pit and recover the ore from the floor of the pit to a depth of 792 m. Development of the underground mine started in 1969. Full production of 273 000 tonnes per month was attained in 1977. The third stage of mining called No.2 Mine, was initiated to accommodate a production expansion and included the development of a second underground mining operation to recover the ore from the bottom of No.1 Mine, i.e. the 2600 level; to a depth as yet undetermined but most likely the 4600 foot level. Development of the No.2 Mine began in 1974. Full production of 91 000 tonnes per month will be achieved in 1982. The choice of the 2800 level as the partition between the two underground mines was based on the following factors: The 2600 level is the lowest level which can conveniently be mined through the No.1 Mine crusher on the 2800 level. Development and trial stoping methods for mining at depth could be evaluated in the 2600 to 2800 block utilizing the existing hoisting facilities of No.1 Mine. Early development of the ore in the 2600 to 2800 block would provide additional mill feed sooner than could be achieved by developing from No.2 Shaft. The underground mining plan then is to mine from the pit floor down to the 2600 foot level and at the same time mine from the 2800 level down to at least the 4600 foot level. PLANNING The Kidd Creek operations are designed to follow a logical mining sequence. The grade of ore supplied to the concentrator is the result of following the mining plan. The philosophy of mining has been to mine downward in horizontal slices. In the case of No.1 Mine the slices are 122 m thick while in No.2 Mine the slices will be 183 m thick. The point of this approach is to mine the pillars immediately after creating them rather than allowing them to stand, deteriorate and generate uncontrollable situations. Two types of ore are handled separately to optimize recoverable values in concentrates. The 'A' type ore is a copper zinc ore with some silver and little lead. The 'C' type ore is a zinc silver ore with some lead and copper

Jan 1, 1981

By Hiromu Kusakawa, Tetsuo Kanematsu

INTRODUCTION In urbanized area with heavy street traffic and with environmental requirements to lessen construction noise and vibration, the cut-and-cover is not always a feasible method to build a small diameter tunnel for sewage, water or other utilities. That is where the use of the shield machine to bore such a small diameter tunnel should be considered as a more acceptable alternative. Although the shield machine of this type should be small enough to meet specific requirements for a given tunnel, it should be big enough to provide some working space inside of the tunnel. Recently, we have successfully developed the fully mechanized small diameter tunneling method which enables us to bore a tunnel as small as 600mm in finished diameter. This paper discusses the concept of this newly developed method, including a case history where sewage tunnels of 800mm finished diameter were constructed by using this method. OUTLINE OF METHOD Developed from the original slurry shield pipe-jacking tunneling method, this method is highly automated so that all operations inside of the tunnel can be remotely controlled. Various data including the advance rate, position and inclination of the shield are transmitted to the central control room outside of the tunnel and monitored by the CCTV and the instrument panel at all times to

Jan 1, 1981

Dragline performance drops considerably when chop-down and double handling of material is needed due to increasing overburden thick- ness or total pit depth. Beyond a certain pit depth the dragline needs support by a pre- stripping system in the upper portion of the overburden. In soft to medium hard ground the bucket wheel excavator can be used for pre- stripping, whereby several methods of material transport are investigated. Conveyor transport around the pit and dumping by stackers is the best solution from the technical point of view. Cross pit transport and dumping by means of a conveying bridge is the method with the lowest costs per volume moved. Cross pit transport by use of a beltwagon with overlong discharge boom is both, poor from the technical point of view and high in overall costs.

Jan 1, 1981

By Clinton A. Hollingsworth

The Flotaire Flotation Cell is a simple, efficient hydraulic/pneumatic Flotation Cell that has no moving parts. It is normally used in conjunction with a surfactant (frother) in which case the combination is known as "The Flotaire System". On many ores, metallurgical efficiencies are substantially increased and operating costs reduce. For example, in floating phosphate rock, it is not unusual for BPL recovery to be increased more than 10%, the major increase being due to floating coarse phosphate, up to 14 mesh, and the difficult to float phosphate. Substantial improvements have also been made on coal, mica, vermiculite, kaolin, barite, copper, etc.

Jan 1, 1981

By M. D. Flavel

Testing methods are described to help in decisions of selecting the most viable comminution circuits. The authors present a comparison evaluation based on numerous studies to help make initial choices for feasibility evaluations. An update summary of completed Bond grindability and Abrasion Index testwork at Allis-Chalmers Process Research Test Center for all materials is included. "What is the best circuit to comminute primary crusher product to make beneficiation feed?" This frequently asked question has the implication that there is one circuit which is universally the right circuit. The answer has been sought ever since it has been necessary to break rock to liberate entrapped mineral values. With every new contribution to the art and the science of comminution, the scope of the discussions broaden. It is not even possible to define what the abstract term best means.

Jan 1, 1981

On December 16, the General Services Administration (GSA) suspended weekly sales of surplus silver from the national defense stockpile because Congress approved the Defense Appropriation Act, which contained an amendment to halt silver sales. The amendment, introduced by Senator James A. McClure (R., Idaho), prohibits sales for 6 months while the impact of the sales and alternative methods of disposal are evaluated. The GSA sold 2 million troy ounces since sales started October 14. 46.5 million troy ounces were authorized to be sold during this fiscal year and an additional 58.6 million troy ounces were authorized to be sold over the following two fiscal years. The stockpile held 139.5 million troy ounces of silver when sales began in October. Brush Wellman, Inc., a vertically integrated producer of beryllium products announced plans to modify its Delta, Utah beryllium concentrate processing facility to utilize beryllium concentrates containing as little as 7% beryllium oxide. The company plans to invest $2.5 million for this modification with completion expected in mid-1983. Implementation of this plan to consume lower grade beryllium minerals will significantly increase world reserves of beryllium concentrates.

Jan 1, 1981

Smoky Valley Mining Company, Round Mountain, Nevada, was the pioneer operation designed for large scale heap leaching of crushed gold ores. It is located in the Big Smoky Valley which stretches 160 km (100 miles) through central Nevada west of the Toiyabe Mountain Range. Gold was first-discovered at Round Mountain, 88 km (55 miles) north of Tonopah, in 1905; and mining in the area included placer operations, which ran from 1946 to about 1960. Smoky Valley Mining Company is a partnership of Felmont Oil Co., Case-Pomeroy & Co., and Copper Range Co., with the latter being wholly owned by Louisiana Land and Exploration Co. In 1969, an option on the Round Mountain property was acquired by Copper Range, and during 1974 a decision was reached to proceed with mine development and plant construction. Ore reserves were reported at 10,537,000 mt (1 1,617,000 st) of 2.09 g per mt (0. 061 oz per ton) gold using a cut- off grade of 0.69 g (0.02 oz). Silver content was approximately 2.4 g per mt (0.07 oz per ton) and the waste to ore ratio was about 1: 1. By the end of 1976 heap leaching was in operation. The ore is a rhyolite tuff, sometimes described as a gold porphyry with highly abrasive qualities but which does not break down into slimes upon contact .with cyanide solutions. Gold is unevenly distributed through the ore but as a whole it is in a very finely divided state. Climatic conditions are extreme in central Nevada. Summers are hot and dry while winters are windy and cold with subfreezing temperatures. Elevation of the property is about 1,920 m (6,300 ft). Precipitation averages 203 mm (8 in.) per year. Until February of 1980, leaching was continued until the temperature reached minus 2ºC (28º F), and about 10 months leaching was the average. Water for the operation is drawn from the Toiyabe Mountains through a 13. 7 km (8. 5 mile) pipeline, originally installed in 1915, or from wells in the valley bottom. The plant requires make-up water at the rate of about 1,500 liters per min (400 gpm). On a schedule of three shifts, seven days per week, the daily output at Round Mountain amounts ,to 12, 700 mt (14, 000 st) consisting of 6,350 mt (7,000 st) each of ore and waste. Ore is hauled with a fleet of 45 mt (50 st) trucks and dumped into the 1,0.70 by 1,650 mm (42 by 65 in. ) gyratory primary crusher located near the pit. The primary

Jan 1, 1981

President Ziolkowski called the meeting to order at 9:30 a.m. He called on the Auditing Committee to give its report. Jack Leech read the report and asked that it be approved. A motion was made, seconded and approved. The Resolution Committee was asked to give its report. The following was submitted: (1) RESOLVED, WHEREAS, in the year since the last convention of The Rocky Mountain Coal Mining Institute, the Institute and the Coal Mining Industry have suffered the loss through death of the services of the following: James A. Flint John L. Paynich Joseph E. Ward NOW, THEREFORE, BE IT RESOLVED, that there be spread upon the minutes of The Rocky Mountain Coal Mining Institute recognition of the passing of these members, who have so greatly contributed to the advancement of the industry, both nationally and in the Rocky Mountain area, and that a copy of this resolution be sent to the next of kin as evidence of the sincere regard and sympathy of the members of the Institute.

Jan 1, 1981

By I. Liebman

The Bureau of Mines conducted experiments on coal dust explosion suppression. The investigation took place on a beltway in the double-entry of the Bureau's Experimental Mine. Using conditions that simulate those in a working mine, the investigation showed that a coal dust explosion can be readily initiated in a conveyor belt entry. Also, the explosion can spread into adjacent entries through opened crosscuts (stopping destroyed by explosion) to propagate for long distances even though the adjacent entries are rock dusted in compliance with regulations. Passive water barriers (water-filled tubs) were found to be effective in suppressing explosions on a beltway; however, barrier efficiency was reduced by opened crosscuts. The study indicated that the beltway barrier should be the distributed type consisting of continuous rows of water-filled tubs covering long distances along the entry. A plan is outlined for the installation of a water barrier system in a working coal mine for the protection of a beltway.

Jan 1, 1981

By Robert B. Haagensen

While light optical microscopy has long been employed by the mineral processing industry, there has been a tendency to shy away from its use for all but the most exacting purposes. Despite the usefulness of micros-copy in solving a wide variety of problems, the response to such work is often less than enthusiastic primarily due to time considerations. Microscopical techniques are therefore often relegated to the "ivory tower" and their direct and practical application to everyday problem solving over-looked. The purpose of this paper is to dispel such attitudes and promote a wider use of light microscopy. Several examples are given which point out the indispensib1e nature of optical microscopy as a rapid means of defining and solving problems in mineral processing. The application of other instrumental techniques to augment the information obtained is also discussed, particularly the value of the combined use of the optical microscope and electron probe micro-analyzer.

Jan 1, 1981

By J. R. Iwinski

Escanaba, Michigan, is located on the South¬ern shore of the Upper Peninsula on a sheltered area of Lake Michigan called Little Bay de Noquet. The Chicago and North Western Transportation Company has operated a rail-to-ship ore transfer facility at this location for well over 100-years. It is the only terminal of its type on Take Michigan and thus can offer a longer navigation season to lower lake markets than other upper lake ports. It all started in 1864 when the Peninsula Railroad of Michigan, now a part of the C&NW, completed a 100 km rail line from Little Bay de Noquet to the Jackson Mines at Negnauee, Michigan. During that first year of operation approximately 31 000t of ore were loaded onto ships at Escanaba from what was then called the Merchants Dock. By 1871 demand had already exceeded capacity and it was necessary to build a larger dock, which in turn was replaced with a still larger pocket dock in 1903. There are no plans or records available to indicate how the ore was handled from rail to ship prior to the construction of the pocket dock.

Jan 1, 1981

By Michael G. Pojar, Peter G. Chamberlain

Solution mining, primarily by heap leaching, has emerged as an important method for mining gold and silver in the United States. Heap leaching offers many economic and environmental advantages over conventional production techniques, in particular for lean ores and old mine waste rock. In order to more effectively pursue research on heap, dump, and in situ solution mining of gold and silver, the Federal Bureau of Mines conducted an extensive survey of leaching practices and problems. This presentation summarizes trends and statistics. Leaching is projected to provide 20 pct of primary mine produced gold and 10 pct of primary silver by the end of 1982. Current leaching practices in the United States are reviewed; operational problems and research programs to alleviate those problems are discussed.

Jan 1, 1981

The Bureau of Mines conducted large-scale laboratory copper ore leaching tests to enable improving copper recovery from large dump-leaching operations. This was done by comparing leaching progress when copper was recovered from recycled leach liquor either by cementation or by solvent extraction. Copper was leached from 7-ton ore samples at room temperature using pH 2 ferric sulfate solutions. With are samples containing 11 percent minus 0.5-inch fines, no marked advantage in using either of the copper recovery methods was evident after 416 days of leaching. With ore containing 32 percent minus 0.5-inch fines, solvent extraction of copper was advantageous in that 60 percent of the copper was extracted after 568 days, whereas only 48 percent of the copper was extracted when cementation was used. Less acid was consumed during leaching when solvent extraction, rather than cementation, was used for copper recovery. Iron, contributed by cementation with scrap iron, plugged the ore bed containing 32 percent fines.

Jan 1, 1981

By L. R. Mabson, F. M. Russell

INRODUCTION Roan Consolidated Mines Limited (RCM) operates five underground mines - Mufulira, Luanshya, Baluba, Chambeshi and Chibuluma (figure 1) . Associated metallurgical plants provide for the operations of milling, smelting two refineries for copper and a cobalt plant. Primarily a producer of high quality copper (242,000 tonnes during 1980) the Company is also the world's second largest producer of cobalt metal. RCM's remaining ore reserves at 31st March 1980 were 275 million tonnes at 2.84% copper. Monthly ore production is in excess of 1.2 million tonnes. Sublevel open stoping was the choice of mining method for the first underground operations on the Copperbelt for ore with dips steeper than 30 degrees. This was for use at the Roan Antelope mine, subsequently to be called Luanshya mine. The sublevel open stoping system was planned based on the methods used in many of the iron mines of Michigan and New York and in the copper mines of Tennessee. Peterson (Ref. 1) gave the following reasons for employing sublevel stoping, reasons which are still valid today: (1) the working place could always be barred and made safe, and the sublevel exit was always at hand; (2) the working conditions were similar day by day and drilling operations were repetitive; (3) pillar spacing can be varied as conditions demand; and (4) stoping operations could be shut down completely or opened up to full capacity at short notice, without affecting grade or causing wastage of reserves. Such was the success of the method that it was adopted as the basic mining method at all the RCM mines in their initial stages. Although cut and fill, block caving, and sub- level caving methods have been used, sublevel open stoping in its many variants still accounts for nearly 90% of total ore production in the Company. The following descriptions outline some of the applications of sublevel open stoping within RCM. LUANSHYA Geology There are two main ore deposits at Luanshya: the Roan (Luanshya Mining Area), in which mining began in 1930, and Baluba, in which mining production started in 1973 (figure 2). Both orebodies occur in synclinal and variably folded basins. These basins are underlain and completely surrounded by much older schists and granites. The ore formation, an argillite, lies near the base of the sediments and is underlain by conglomerates, quartzites and arkoses. Copper mineralisation is mainly found in two well defined zones, separated by a varying thickness of pyrite mineralisation. In the eastern part of the Roan deposit only the upper orebody is present, but west of 18 shaft and at Baluba the lower orebody predominates. Locally the subjacent quartzites of the footwall formation are also mineralised with copper. The copper mineralisation is found princi- pally as disseminated sulphides, mainly chalcocite, bornite and chalcopyrite. Near the surface the sulphides have been changed to oxides. At Baluba the cobalt mineral carrollite is found with the sulphides. Mining methods Sublevel open stoping is the basic mining method. In its standard form this method is best suited to steeply inclined orebodies, with simple structures and with a strong foot- wall and hangingwall. In recent years, as more complex folding and flatter dipping orebodies have been encountered, considerable variations have been introduced. Stope drilling is effected with drifters and extension steel.

Jan 1, 1981

By John R. Alderman

INTRODUCTION Uranium production in Saskatchewan commenced in 1953 at the Eldorado Nuclear Mine in the Beaverlodge area of North Saskatchewan, followed over the next few years by nineteen other smaller mining operations. Production rose from 453 600 to 2 721 600 kg per annum by 1958. Over supply and the closure of the U.S. market to Canadian uranium led to an equally rapid decline to an annual production which varied from 453 600 to 907 200 kg during the years 1964 to 1976. As the market strengthened there was a resurgence of the Saskatchewan industry and production from one underground and two open pit mines reached 2 809 000 kg in 1980 for a return of 232 000 000 dollars (Table 1). [ ] Future output will be increased by a potential 4 536 000 kg from the Key Lake mine now under construction and scheduled to commence production in 1983. Three other projects are in an advanced stage of planning and are currently in the regulatory review process. Several other prospects are being evaluated. Figure 1 shows the location of uranium mines in Saskatchewan and table 2 illustrates the present status of uranium mines and deposits in Saskatchewan. At present, 1400 workers are employed directly by the uranium mining industry. Saskatchewan has diverse mining conditions ranging from the underground mine of Eldorado Nuclear Limited 1870 m deep with a low grade ore deposit averaging from 0.17-0.20% with extraction by carbonate leaching through an open pit operated by Gulf Minerals with a moderate depth of 110 m and 0.37-0.5% grade to the shallow open pit of Cluff Mining 30 m deep with a very high average ore grade of 7.0%. Extraction at both open pits utilizes acid leach extraction processes, the latter after gravimetric separation. Key Lake mining has commenced development of the next mine to come into production which will have two open pits to a depth of 80 and 175 m with an average 3% grade to be followed by the deep open pit of Canada Wide Mines Ltd. to a depth of 205 m with average grade of 1.5%. One existing mine will transfer its operation from the present open pit to another of slightly higher average grade. From the Saskatchewan point of view, these different operating conditions present varied problems associated with occupational health and safety and concomitant solutions to reducing worker exposure to radiation. REGULATION OF URANIUM MINES In 1946, the Government of Canada enacted the Atomic Energy Control Act which established the Atomic Energy Control Board (AECB) and vested in that board powers to regulate all aspects of uranium mining. In the early years, the AECB relied heavily on provincial occupational health and safety agencies for the surveillance of uranium mining and milling operations, particularly with regard to "conventional" occupational health and safety, the AECB having at the time no specific regulations covering the mining industry. There has been considerable dialogue between federal and provincial governments since, traditionally, occupational health and safety is a provincial jurisdiction. Saskatchewan is strongly opposed to the intrusion of the federal government into the provincial domain, particularly as there is little justification for a federal presence in occupational health and safety in uranium mining and milling. The province does, however, accept the need for federal involvement in certain aspects of the regulation of nuclear industry, for example, international sales, research and security of reserves for domestic use.

Jan 1, 1981

By S. K. Bose, P. K. Sahoo, S. C. Sircar, P. K. Jena

High grade manganese ores are utilized in metallurgical, chemical, and battery industries, whereas the low grade ores are not accepted for these purposes and hence get accumulated at the minehead. In the present study an attempt was made to solubilize the manganese value as its sulfate by roasting the low grade ore with a suitable sulfatizing agent followed by water leaching. The different parameters were determined in the laboratory scale. Based on these, kilogram scale experiments were conducted where an appreciable amount of manganese goes into solution during leaching.

Jan 1, 1981

By Cudmore JF, Rowlands D

Incidents associated with spontaneous combustion of coal have resulted in the loss of many lives and considerable capital and equip- ment. Nevertheless, the factors controlling the propensity of coal to spontaneous combustion are not clearly understood and there is a need for research to study the effects of rank, petrog- raphy, moisture contents, etc. This has been attempted in a study of the propensities of selected Queensland coals to spontaneous combustion. An adiabatic calori- meter was constructed to study the self-heating of a large number of coals. Chemical and petro- graphic analyses of the coals were determined and an attempt was made to relate these propert- ies to the self-heating characteristics of the coal. The results show a strong relationship between rank and self-heating characteristics and that petrographic composition of coal may also be of importance. Some observations on the application of the laboratory results to the practical situation are made as are recommendations for future research.

Jan 1, 1981

By Upfold RW, Velzeboer MD

An underground mine fire will influence the ventilation equilibrium in several ways which results in a redistribution of the ventilating quantities and pressures throughout the whole mine. The ventilation carries the combustion products which constitute the greatest danger of mine fires through the mine and provides the oxygen on which the intensity of the fire depends. However, as a result of the ventilat- ion disturbances created by the fire, changes in airflow patterns can take place which are not always obvious, especially when the mine is being served by several fans. The use of com- puter simulation techniques allow for the likely effects to be tested and the different ventilation options to be evaluated as an integral part of the overall emergency/rescue planning and control procedures.

Jan 1, 1981

By S. R. Crane

One Bureau of Mines goal is to minimize the undesirable environmental impacts associated with industrial plants emitting waste gases containing S02. To help meet this goal, a regenerable flue gas desulfurization process was developed. This process, known as the citrate process, uses a buffered weak acid solution to absorb S02 from the waste gas. The absorbed S02 is reacted with H2S to precipitate elemental sulfur and regenerate the solution for recycle. The H2S feedstock for the process, if not otherwise available, may be produced by reacting two-thirds of the recovered elemental sulfur with natural gas and steam. Laboratory investigations and pilot plant operations were conducted by the Bureau to determine if H2S from the natural gas, sulfur, and steam reaction was suitable for the citrate process. The laboratory investigations, in which an H2S generator was integrated with other citrate process operations, provided a basis for design and operation of the pilot plant. The objective of the pilot plant was primarily to provide H2S for the citrate pro-cess pilot plant at the Bunker Hill Co. lead smelter in Kellogg, Idaho. The design capacity of the pilot plant was 0.4 to 1.25 tons of H2S per day. The operation demonstrated that (1) a 77-to 79-dry-volume-percent H2S product could be produced from pilot-plant product sulfur, natural gas, and steam; (2) the product R2S regenerated the SO2-loaded citrate solution; and (3) under certain conditions over 95 percent of the natural gas could be converted to H2S.

Jan 1, 1981