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By Hans Ehdwall

INTRODUCTION Investigations of radon and radon daughter concentrations in Swedish [non-uranium] mines started in the late 1960's. The first screening measurements showed that the average annual exposure to radon and radon daughter products was 4.7 WLM. The main reason for high radon and radon daughter concentrations was inefficient ventilation and radonrich water entering the mine. In the radon regulations worked out later it was stated that no miner should be exposed to more than 60 000 pCi h/1 equilibrium equivalent concentration of radon annual exposure, corresponding to 3.6 WLM. Now, 1981 the situation has changed considerably. From the average annual exposure of 4.7 WLM in 1970 it is now only 0.7 WLM. Sweden has up to now had only one [uranium] mine and the work there has only been investigative. However, there are plans for a commercial uranium mine in another part of Sweden. The radon problems in these mines are widely different depending on the mineralogy. NON-URANIUM MINES The radiation problems in Swedish mines were not recognised until the late 60's. The first radon and radon daughter measurements were made in some sulphide ore mines in 1967 (1). The radon and radon daughter concentrations were surprisingly high for non-uranium mines. In order to have a complete picture of the radon situation in Swedish mines the National Institute of Radiation Protection (NIRP) decided to make measurements in all, at that time about 60 mines (2). To get results as fast as possible measurements on radon gas seemed most appropriate to start with. Sampling was made by mailing a number of evacuated 4.8 litre conventional propane containers from NIRP to each mine. The containers were then opened at the place of interest. After sampling the containers were sealed and then mailed back to the institute for measurement. The measurements were made in ionization chambers. This method only gave the radon concentration and the radon daughter concentration was estimated by multiplying the radon concentration by an assumed equilibrium factor. The equilibrium factor is defined as the ratio of the total potential alpha energy for the given daughter concentration to the total potential alpha energy of the daughters if they are in equilibrium with the given radon concentration. The results of this first preliminary survey indicated that a great many of the Swedish miners probably had an annual radon daughter exposure of more than 3.6 WLM. As the radiation exposure in non-uranium mines was not regulated in either the Swedish Radiation Protection Act or the Swedish Labour Protection Act work was started on special radon regulations. A lung cancer mortality study was also started. To check the results of the first survey and to get experience and knowledge of radon problems in mines, it was decided that personnel from the NIRP should visit each mine for a detailed investigation of radon and radon daughter concentrations starting with the ones with the highest radon concentrations. The main reasons for these so-called "basic measurements" were: 1. To estimate the doses received by Swedish miners 2. To find the sources of the high radon and radon daughter concentrations 3. To find appropriate counter-measures 4. To determine the most typical equilibrium factor for each mine. Unlike most uranium mines the reason for high radon concentrations in non-uranium mines is seldom the occurrence of highly radioactive minerals. The main sources were found to be waste-rock and radon-rich water. In order to filter and warm up the inlet air, especially in winter time, it was very common at that time to suck the air through broken wasterock. By doing so the air was contaminated with radon from the waste-rock and radon-rich water in it. It is noteworthy that the radium and uranium concentration in the waste-rock is relatively low. The uranium concentration is only of the order of 15 - 20 ppm. The action to prevent this contamination of the inlet air was to change the direction of the ventilation and in the case of radon-rich water entering the mine the action was to prevent the air coming into contact with the water. The first calculation of the radon daughter exposure of Swedish miners was based on radon gas measurements. The radon daughter concentration was estimated by using an assumed equilibrium factor of 0.5. Later when the mines were visited by institute staff it was possible to compare the assumed equilibrium factor with the measured ones. It was found that the factor varied from 0.15 at the air inlet to 1.0 at the air outlet and the average equilibrium factor on workplaces for almost all mines was between 0.4 and 0.6. The result of the exposure calculation in 1970 showed that more than 40 % of the miners had an annual radon daughter exposure of more than 3.6 WLM. The overall average was 4.7 WLM and the maximum annual expo-

Jan 1, 1981

By Ireneusz Tomza, Jolanta Lebecka

INTRODUCTION Radioactive deposits were observed in 1972 in some of the Upper Silesian coal mines. They were located mainly in the drains in galeries and on the inside surfaces of water pipes. They also caused some problems by accumulating in water pumps. It has been postulated that the deposits are produced by natural radioactive waters seeping from the rocks. Investigations were initiated to answer the following questions: - What is the composition and the amount of radioactivity in the deposits? - What radioisotopes are present in the water? - How are the radioactive deposits formed? - Do the radioactive waters also occur in other mines? - How does the radioactivity of the water depend on chemical composition? - What is the origin of the radioactive water? - Does the water and the deposits cause radiation hazards for miners? -How can the radiation hazard be reduced? METHODS OF MEASUREMENT Determination of Radium Isotopes in Water The commonly used methods of radium determination in water are either based on measurements of the radioactivity of 222Rn which is in equilibrium with 226Ra, or on the detection of alpha particles of the radium radioisotopes after chemical separation of radium from the water sample. The method based on radon activity measurements is very sensitive and does not require any chemical Separation, but it can be used for determination of 226Ra from the uranium series only, because the thorium daughter 220Rn has too short a half-life (55s to yield the required accuracy. The method developed by Goldin, 1961 [2] involved alpha-particle measurements in thin layers of RaS04 and BaS04 separated from the water. This method is not convenient for saline water and water with high barium concentration because the amount of barium carrier in this case is too large to obtain a thin layer of precipitate with sufficient activity. The Upper Silesian carboniferous waters are often saline with high barium content, so the method described by Goldin was not convenient for this case and it was necessary to change the detection system and modify the chemical preparation. The procedure developed by the authors for the determination of radium isotopes in water was as follows: - Depending on the Ba2+ content and the required sensitivity of measurement, a water sample of 200 cm3 to 3 dm3 was taken. - 10 cm3 of 0.25 M citric acid and 5 cm3 15M ammonia was added to form complex Ba2+ ions and avoid the immediate precipitation of BaSO4. (This was repeated as long as the addition of BaC12 did not form a precipitate.) - 1 cm3 of 1N solution of Pb(N03)2 as a carrier for radioactive isotopes of lead and 10 cm3 of 0.1 N BaC12 as a carrier for radium were added. - The sample was heated to the boiling point and the precipitation of RaS04, BaSO4 and PbS04 with 50% H2SO4 was carried out. - After several hours the sample was centrifuged and the precipitate was purified by washing with nitric acid and distilled water. - The precipitate was then redisolved in 20 cm3 0.125 M Na2EDTA and 3 cm3 6M ammonia and reprecipitated from the solution by dropwise addition of acetic acid to d pH of 4.5. At this value of pH, precipitation occurs only for the barium and radium sulfates, while lead and all other radioactive elements remain in the solution. The date and time of deposit precipitation was recorded. - The final barium-radium sulfate mixture was washed with distilled water and transferred to standard measurement vials. - Each vial containing a deposit had 6 cm3 of distilled water added and was then shaken vigorously. 12 cm3 of liquid gelling scintillator (INSTA-GEL UNISOLV-1 type) was then added and the vi 1 was shaken again. After a while the scintillator turns into a milky gel in which the deposit is uniformly distributed. - The standard sample of 226Ra was prepared in the same way. - The activity of the samples was measured using a liquid scintillation spectrometer. (In this case the TRICARB 3320 produced by Packard Instruments, was used). Tests run on standard radium solutions provided by Amersham Radiochemical Centre indicated that this method of measurement enables one to achieve an efficiency of almost 100% (within measurement error). For alpha particles no quenching effect was observed for the BaS04 concentration in the range up to 80 mg of BaS04 per 1 cm3 of liquid scintillator coctail (Fig. 1). This provides a sensitive determination of radium in water with high barium content and also in saline water. In saline water the solubility of barium sulfate is much higher than in

Jan 1, 1981

By Jon C. Volkwein

The U. S. Bureau of Mines investigated a method of using inert gas to prevent the formation of explosive gas mixtures in auger highwall mining of coal. A combination of gasoline and diesel engine exhaust gases was introduced into the auger drill hole using a short section of pipe located at the collar. Gas samples were taken and analyzed on site with infrared detectors for oxygen, carbon dioxide, methane, and carbon monoxide. Evacuated bottle samples were also taken and analyzed by gas chromatography at the Pittsburgh Research Center. These gas results were analyzed for explosibility. Personal exposure to carbon monoxide was also monitored. The highest methane level observed was 9.55 pct. The Inert gas levels, (carbon dioxide and nitrogen) were sufficiently high to prevent any ignition of the methane. Results showed that for all conditions during mining, gas concentrations were non-explosive. The maximum personal time weighted average sample for carbon monoxide was 20 ppm. This system provides a safe, inexpensive, simple method for preventing explosions during auger mining. INTRODUCTION The auger highwall mining method is an effective method to recover coal from a reserve when removal of the overburden by surface mining equipment becomes uneconomical. In this method of mining, a horizontal auger enters the coal seam from the surface mine bench under the highwall and the coal is drilled in a series of parallel holes. Historically, coal mined from the surface is relatively shallow, and over time, methane associated with the coal has dissipated through the surface. In most circumstances, little methane has been found associated with auger mining. However, mining technology has enabled surface mining of deeper reserves of coal. Furthermore, environmental constraints have forced the highwall extraction method to be used to remove coal under wetlands, further increasing the chances of encountering methane. Recently incidents of methane explosions at a few auger mining operations have resulted in injuries and increased testing for methane at the collars of auger holes. The fuel source of the reported explosions was not necessarily limited to methane, but may also have involved coal dust. The Mine Safety and Health Administration (MSHA) met with the Bureau to discuss what technology might be available to enable the safe resumption of mining. The discussion included the difficulty of ventilating through the solid shafts of the augers, that steel bits probably created the ignition source, and that perhaps inerting the holes with low oxygen and high carbon dioxide concentrations from the machine's diesel exhaust was a potential solution. Considering the ventilation aspects of the problem, it was not clear If ventilation could be reliably established. If some degree of ventilation to the front of the mining head is achieved, it may combine with methane to bring the hole atmosphere from a rich, nonexplosive mixture to an explosive mixture. Furthermore, it may not prevent a dust explosion in such a mining configuration. Lack of access through the shafts of auger type mining machines further limits the ability to add water or air to cool bits to prevent an ignition source from developing. Either of these approaches would also be expensive. The process of mining coal In an inert atmosphere has been considered in the past, but to our knowledge, never implemented (Department of Interior, 1970). Clearly, implementation in underground mining would be more complicated. On a mine bench open to the atmosphere, however, adding inert gas to the mining head could provide a quick, feasible method to prevent explosions at auger highwall mining operations. Also the problem of how to move the inert gas to the cutting head of the machine had to be considered. Preventing explosions on auger mining machines using inert gas requires three primary considerations: first is the source of inert gas; second, placing the inert gas at the cutter head; and third, monitoring the hole atmosphere. Any gas source having an effective inert gas concentration of 34 volume pct or greater will prevent methane from Igniting (Zabetakis, 1965). Sources of inert gas considered for this application included liquid nitrogen, modified shipboard inert gas generators (for hydrocarbon shipping and transfer), jet turbine engine (Paczkowski, et. al., 1982), the auger's diesel engine and a gasoline engine. Operation cost, purchase cost and availability limited our testing to the diesel and gasoline engines. This work tested each engine, separately and combined. To ensure effectiveness, both company and enforcement personnel need to know how to monitor the condition of the inerted hole. Measurements at depth inside the hole are possible by remote sampling through rigid tubing, but this method is Impractical for routine monitoring. Continuous monitoring of the exhaust gas stream is an alternative. The U. S. Bureau of Mines evaluated an inert gas system at an auger mining operation at a surface mine near Owensboro, KY. Coal was mined from the Number 9 Coalbed in Henderson Co. KY. Tests were conducted in January and March of 1992.

Jan 1, 1993

Last year in the US alone, about 425 Mt (468 million st) of minerals and coal were beneficiated by froth flotation. This number indicates that from 1983 there was a 10% increase in tonnage of min¬erals and coal beneficiated by the indus¬try. A significant improvement was seen in the tonnage processed by the nonferrous minerals and coal industries. BP Minerals America installed 85 m; (3000 cu ft) flotation cells at the Bing¬ham Canyon mine and concentrator. The new flotation circuit has fewer than 100 cells compared to 2000 flotation cells used in the old plant (Mining Engi¬neering, November 1988). Column flotation use on a commer¬cial scale continues to expand as seen from the interest expressed at the Col¬umn Flotation Symposium (Column Flotation '88). The Magma Copper Co., San Manuel Division replaced all con¬ventional cells with 1.8 x 12 m (6 x 40 ft) column flotation cells for copper con¬centrate cleaning. Also, 1220 mm and 760 mm-diam (48 in. and 30 in.-diam) column cells are operating at the plant in the molybdenum circuit. A commercial Diester Flotaire col¬umn cell for fine coal recovery was installed at the United Coal Wellmore No. 20 plant. The 36.8 m3 (1300 cu ft) cell recovers 13.6 to 18 t/h (15 to 20 stph) of -590 gm (-28 mesh) coal. A similar unit has been installed at Tanoma Mining Co. in Pennsylvania. Various modifications of the column cells are being designed around the world. Jameson (Mining and Metal¬lurgy, 1988) described a new concept whereby the feed and air stream mixture is discharged into a cylindrical column of about 1.2 m (4 ft) height. Recovery and grade of nonferrous minerals have been reported to be better than that in a four-stage conventional flotation clean¬ing circuit. Flotation reagents American Cyanamid and Dow Chemical continued development of a new generation of sulfide collectors. A general feeling is development of new sulfide collectors has not kept up with flotation technology. Additionally, joint efforts between industry and chemical suppliers will likely be necessary to realize the economic benefits of the new technologies, since new chemistries respond differently compared to the conventional collectors. Flocculant development in recent years has been evolutionary rather than revolutionary. Rothenborg reported on development of a new flocculant family (a hydroxymated polyacrylamide desig¬nated S-6703) that has shown consider¬able promise in red mud clarification. Plant testing showed that this new floc¬culant could replace starch and poly¬acrylate and provide significantly higher overflow clarity. Barol Kami (Siirak) and Cleveland¬Cliffs (Hancock) reported development of an amphoteric apatite collector (ATRAC 873) that was used in Tilden's silica flotation process to increase apatite rejection. The collector was engineered for the particular flotation conditions in the complex Tilden process. Significant plant testing with ATRAC 873 showed that this reagent gave significantly in¬creased apatite rejection without any effect on silica flotation effectiveness or selectivity. Electrostatic separation Electrostatic separation is now em¬ployed in the precious metals mining industry to recover gold and silver grills from crushed slag. The installation at Paradise Peak has prompted other op¬erators to consider this application. In another development, attractive potentials for treating very fine minerals (-45 µm or -325 mesh) are being devel¬oped by Advanced Energy Dynamics and by the Department of Energy. Demonstration tests using triboelectric charging/electrostatic separation have been successful on a variety of minerals as well as coal. Magnetic separation Developments in magnetic separa¬tion have transpired on a production scale. Superconducting, high gradient magnetic separation has gained accep¬tance with the successful startup of a second unit treating kaolin at J.M. Huber Corp. This liquid-helium-cooled mag¬net generates 2.0 tesla in a 3-m-diam (120-in.-diam) bore with no power con¬sumption. Wet, high-intensity magnetic separation has been applied to sulfide mineral separations both domestically and abroad. These continuous type of separators are effective in removing residual chalcopyrite and sphalerite from other base metal sulfide concentrates. Separators using high energy rare earth permanent magnets are continu¬ally increasing. Now offered as both drum and roll type, these units are be¬coming a staple in the processing of industrial minerals. Tests using rare earth magnets strategically placed on a spiral concentrator have demonstrated the enhanced recovery of heavy miner¬als such as ilmenite. Classification Although no major technology break¬throughs in classification appear immi¬nent, there is an increasing need for more efficient and cost-effective meth¬ods to make size separations. It is be¬coming more apparent that mineral concentration methods will be more common at very fine sizes, say below 50

Jan 1, 1989

By William A. Rhoades

INTRODUCTION Ever since miners were faced with the task of moving ore, some form of scraper has been in use. At first, men and beasts of burden supplied the power to move the scraper, and later, machines were developed for this purpose. In the early 1900s, a few mining properties used small pneumatic single-drum winches to pull loaded scrapers to a raise or ore pocket, and the empty scraper then was dragged back to the muck pile by a miner, as shown in Fig. 1. Just prior to 1920, an improvement was introduced, using two single-drum hoists. As illustrated by Fig. 2, the second hoist was used to return the empty scraper to the muck pile. However, this arrangement still re¬quired two men, one operating each hoist. The next developmental step was to eliminate the second man by locating the hoists side-by-side and using one man to control both hoists. As illustrated by Fig. 3, this involved the use of a tail rope over a sheave to pull the scraper back. The greatest progress in development and the great¬est increase in the use of slusher haulers occurred between 1920 and 1930. In 1921, the Sullivan Machinery Co. designed and built the first two-drum scraper oper¬ating on the principle illustrated by Fig. 4. It was powered by a 4.5-kW (6-hp) Turbinair(r) motor and would pull a 450-kg (1000-1b) load at 0.61 m/s (120 fpm). In 1922, this unit was shipped to the Verona Mining Co. of Caspian, MI, and it experienced immediate success in the Lake Superior iron ore district. Since the two-drum slusher was much less expensive and more efficient than hand mucking, the Lake Superior mines were saved from financial disaster when iron-ore prices fell 25% between 1923 and 1925. Immediately there¬after, a demand developed for slushers that would oper¬ate with electric power, which was considerably cheaper than compressed-air power. In 1923, the Sullivan Machinery Co. responded with the first electric-powered double-drum hoist. During subsequent years, design improvements in¬cluded separate tail-drum gearing to increase the tail¬drum speed, as welt as a number of safety features such as rope guides. One result of these improvements and their utilization by the Michigan iron mines was an in¬crease of 100% in the tons of ore per miner per day in those mines between 1924 and 1929. Since the two-drum slusher was capable only of straight-line mucking, it was not a practical machine for use in open stopes. In 1929, the Sullivan Machinery Co. introduced the first three-drum slusher. As illustrated by Fig. 5, two tail drums and one hauling drum were provided. A tail sheave could be placed at each side of the stope, and the ore then could be loaded and hauled to a central point from the entire width of the stope. During the 1930s, progressively larger slushers were demanded. By 1940, two- and three-drum units were available with motor power as high as 45 kW (60 hp), and slusher power continued to increase after 1940. Be¬tween 1951 and 1952, Joy Manufacturing Co. designed a 112-kW (150-hp) two-drum Blusher for the Climax Molydenum Co. Although there has been no demand for a slusher more powerful than this, 150- to 225-kW (200- to 300-hp) slushers are quite feasible at the pres¬ent time. During the last 30 years, many slusher improvements have been made to the operating life, operational safety, and ease of operation and maintenance. Increased tail-drum speeds have decreased overall scraping times. Rope guards, totally enclosed drums, and operator shields have reduced the hazards of injury due to wire¬rope breakage. Improvements in lubrication have made the slushers relatively maintenance-free, with long operating lives. The introduction of spring-actuated drag brakes prevented uncontrolled unreeling of dis¬engaged drums, allowing the development of practical remote-control slusher operation. Remote control now is available in a choice of all-air, all-electric, or air¬electric slushers. APPLICATIONS Quite simply, slushers are used to load and transport material (ore), generally over a short distance of from

Jan 1, 1982

By C. D. Mann

INTRODUCTION The sublevel stoping mining method is usually applied to a relatively steeply dipping, competent ore body, surrounded by competent wall rock. Ore is produced by drilling and blasting longholes, which can range from 50 mm (2 in.) to 200 mm (7% in.) diam, with lengths up to 90 m (300 ft). Longholes can be inclined in any direction, but the ring or pattern usually forms a plane, and the holes are blasted as a unit. Recently developed mobile drilling and loading machinery, as well as new explosives products, blasting techniques, and cemented sand and rock fill have made sublevel stoping a highly efficient and versatile mining method. When designing a sublevel stoping production sys- tem, it should be kept in mind that production rates from conventional sublevel stopes vary widely through- out the life of the stope. Early production is at a low rate, coming only from the drawpoints near the slot, but increases as new drawpoints are reached by the stope face. As the stope nears completion, again, fewer drawpoints are productive. Enough drawpoints must be available at any time to provide required production. Drawpoint availability should be compared to equipment availability; plan for more drawpoints than are needed at any one time. Accurate, realistic scheduling is essential to smooth production rates. Also, initial recovery of ore in a stope/pillar block is normally from 35% to 50% in sublevel stoping. Planning of pillar recovery, representing the majority of ore tonnage in a production block, must be done during early mine planning. Since much of the development already done for primary stoping (access for drilling, drawpoints, and haulageways), can be used for pillar recovery, early production from pillars is highly desirable. The following description of components of the system is an attempt to highlight some of the most important features and requirements of mechanized sublevel stoping methods. Similar comments would apply to the use of older equipment (column-and-arm drill setups, slushers, etc.) in similar methods. As in any good mining system, maximum economic recovery of the resource in the ground is the primary consideration. STOPE DESIGN CHARACTERISTICS Length and Width The following are some of the factors which affect sublevel open stope length and width dimensions: ore body geometry, principal stress directions, competence of stope back, optimum drill pattern, and drilling drift layout. In new mines initial stope layout design may occur before the ore body is actually intersected by mine workings. Stope dimensioning is a critical decision, and assistance from as many knowledgeable people as possible at this stage is essential. Operators with past experience in similar ore bodies, rock mechanics experts, and others with mine design experience should participate at this stage of stope planning. Height The following are some of the factors which must be considered in determining stope height: competence of stope pillar and stope/fill walls; slenderness ratio of adjacent pillars; ore body dip; ore body thickness; hole depth capability of the drilling machine; fragmentation characteristics of the ore; and level intervals in existing mines. In competent ground, drill-hole length and accuracy are the most important determinants of stoping height. Frequently entire drilling sublevels can be eliminated because of the depth capability of sophisticated drilling equipment, resulting in significant development cost savings. Drawpoint Location and Design Some of the most important considerations of a good drawpoint system are optimum spacing of draw- points, within the constraints of stope dimensions, for uniform drawdown and maximum recovery; excavations designed for stability for the life of the ore block to be drawn-primary stope ore as well as subsequent pillar ore; floor or roadway design including type of surface, reinforcing, grade for water runoff; orientation with respect to the main haulageway, for optimum loader maneuverability and ground stability at the inter- section; and length, to allow articulated front-end loaders to work in a straight configuration. Careful drawpoint design and construction are keys to successful production. Extra care in development, such as smooth wall blasting, rockbolts or grouted rebar, wire mesh, and shotcrete usually will ensure long draw- point life. Human exposure during production loading is of longer duration than during development or production drilling, and consequently preparation of draw- points is easily justified, particularly when pillar ore can be drawn through the same drawpoints. Secondary blasting of boulders can weaken drawpoints, also justifying good ground control techniques. A smooth draw- point floor of poured, reinforced concrete, on a grade of +3% or +4% toward the ore pile facilitates water flow out of the drawpoint, and ease of loader bucket penetration into the muck pile. Slot Raising, Slotting A slot or other space for rock expansion is necessary in conventional sublevel stoping where vertical rings or rows of holes are blasted. The slot can be started at a slot raise driven by conventional raising methods, raise boring, drop raising (predrilling and blasting a raise from the top, using small diameter-less than 200-mm (7%-in.)-holes for relief), or crater blasting (similar to drop raising, but without relief holes). The slot usually extends from the extraction level to the back of the stope. It is normally expanded to full stope width by

Jan 1, 1982

By A. Dale Wunderlich

With the price of precious metals at an 18-year low, every ounce of metal produced is important. The theft of metals from mining and refining sites can mean the diffrence between profit and loss for many mining companies. Low metal prices do not reduce the potential for the theft of precious metals. History has shown that the price of gold has little to do with the desire for employees to steal precious or base metals. There is actually evidence that the theft of precious metals increases when the price of this commodity goes down. Several of the major precious metal thefts in the past year took place at silver mines when the price of silver was less than 16 cents/g ($5/oz). How does the lowest gold price in 18 years affect the need for security at precious metals properties? There is no short answer to this question. One reason is because the exposure to theft of precious metals is unique to each property. This makes it important that each property be evaluated individually. More than 95% of all precious metals thefts can be attributed to those working at the mine site. So preventing employee theft is the primary concern. One consideration is the location of the property. Gold selling at any price is still an attractive commodity in countries where the employees are making between US$400 and US$600 a month. It is not uncommon for employees at mines in countries where low wages are the norm to consider the value of a gram or two of gold to be a significant amount of money. A gram or two of gold a day may not seem like much. But if 15 employees steal two grams a day, that equates to a significant amount of money during a year. The type of property where the precious metals product is being recovered is also important. For example, a property with a gravity circuit is more likely to suffer from the theft of gold product than a property where all gold is finely disseminated and the only gold seen in the ore body is through a microscope. Gravity circuits increase an operation's exposure to theft because the grinding circuit that is associated with a gravity circuit often becomes a giant concentrator. Areas such as the bottom of grinding-mill pump boxes, cyclone-feed-pump clean out traps and the sumps often become locations where precious metals concentrate (Figs. 1 and 2). Muck concentrations in these locations can be as high as 25% to 40% of gold or silver. Not long ago, muck was removed from a barren-solution sump at a Merrill Crowe circuit that had concentrated to more than 40% gold. At a milling site in the Pacific Rim, residents of the community adjacent to the mine learned about the value of the concentrates in the sump under the ball mill and committed an armed rob¬bery. While several of their co-conspirators held the em¬ployees at bay with machetes, the others emptied the contents of the sump into buckets and removed it from the site. Armed robbery is not as common as employee theft. However, while this article was being written, an armed robbery occurred at a gold property in Central America. Armed perpetrators took as hostages the night shift employees at a process plant and used cutting torches that were on site to cut into the high-security and gold-storage areas. The perpetrators then stole a company vehicle to remove the stolen gold buttons and sludge from the site. Unfortunately, this type of activity goes on regularly. But managements of most mining companies are reluctant to discuss theft scenarios. So information pertaining to the theft of precious metals seldom becomes a newsworthy item. An audit conducted at a mine site with a gravity circuit recommended that the gravity recovery area be shut down until adequate protection could be provided. Although it was not connected with the audit, it was necessary to shut down the gravity area for a pro¬longed period because of problems with the gravity table. In the two months that followed, gold production at the site increased by about 31 kg/month (1,000 oz/month). It is difficult to attribute all of this increase to the theft of concentrates. But there was a good chance that at least part of the increase was due to the fact that concentrates were being stolen from the gravity area.

Jan 1, 1998

By Francis S. Kendorski

INTRODUCTION Drift design problems in caving operations are a re¬sult of the geologic factors contributing to the overall success of the system, of the engineering factors dictated by economic and technical considerations, and of ore production practices. Combining these factors, a rational underground sup¬port system of rock reinforcement, light steel channel section or welded wire fabric, and shotcrete can be de¬signed based on rock fracturing, rock load, abutment loadings, ground movement, expected repair and desired flexibility. The design concept uses the effect achieved by restraining, reinforcing, and maintaining some of the intrinsic strength of the fractured rock mass composed of interlocked blocks of intact rock and rock fractures. Three different examples of drift support design in hypo¬thetical mines using the caving system are given. Caving is a system of underground mining where ore is extracted by means of gravity after the ore body is allowed to fail by removing support from underneath. The rock mass of the ore body fractures and flows ver¬tically downward to let gravity do as much work as pos¬sible. Caving differs from many other mining systems in that blasting is used only to initiate the rock mass failure by removing the rock supporting the ore but not to break the ore itself. The initial movement of the rock mass dur¬ing failure and the consequent crushing and grinding during the continued movement serve to reduce the ore to particles of a manageable size, with only limited sec¬ondary blasting necessary. The broken ore is extracted from the bottom of the failed rock mass through funnels of some sort pre-excavated in the rock. Ore extraction must continue or the swell of the broken rock will even¬tually fill the cavity and stop further rock mass failure and movement. The excellent general discussion on block caving in the SME Mining Engineering Handbook (Julin and Tobie, 1973) adequately covers the principles and application of this type of underground mining. Many rock mechanics aspects of block caving have been covered by others (McMahon and Kendrick, 1969; Swaisgood, et al., 1972; Mahtab and Dixon, 1975; King, 1946) and will not be reviewed further. Maintaining the stability of production drifts is one of the most troublesome problems plaguing the mine manager in a caving operation. Many factors contrib¬ute to drift support problems, and identifying the causes of instability and producing a reasonable support design are two steps toward achieving stability consistent with the mine plan. This chapter sets forth a technique for the design of support systems for production drifts in caving opera¬tions. The basic support system elements employed are rock reinforcement, welded wire fabric, and shotcrete. Recognized as contributing to the design are the factors of rock load, additional load from mining activity, rock fracture characteristics, repair expected, and flexibility. It must be emphasized that the drift must first be stabilized as for a tunnel, and the additional strengthen¬ing for mining-induced loads cannot contribute to the initial premining stabilization, or the reserve of strength is used up. MINE PLANNING The efficient mine planning engineer not only must satisfy the economic, human, and environmental aspects of his task but must also consider the mechanical con¬sequences of his plan. The problems created for the mine by placing parallel drifts too close, by crossing drifts on different levels with inadequate, if any, separa¬tion, and by installing connections and crossovers in the haulage plan without regard for the effective spans cre¬ated, are only a few of the problems a mine planning engineer can create for himself and the mine. The effect on immediately adjacent mine areas when an area is caved is important to drift design because the removal of vertical support from a rock mass causes the weight of that rock mass to be shifted elsewhere. The adjacent rock mass will carry this load and reach a new equilibrium with the applied stress. The advancing front of stress increase that results from caving (and many other mining systems) is generally called the abutment load and is the increase in stress over the gravity or tec¬tonic stress that already exists, as shown in Fig. 1. In general, the abutment load will be similar in nature to the stress change found around an opening in rock and will be taken as causing an increase in the vertical prin¬cipal stress, due to an approaching caving boundary, so that

Jan 1, 1982

By S. E. Phillipson

The MSHA Technical Support, Roof Control Division's primary evaluation tool for coal pillar designs is the Analysis of Coal Pillar Stability (ACPS) program. However, ACPS is limited to evaluating interactions between two seams at a time, which is not ideal for complex multiple- seam scenarios. Further, it does not recognize the influence of strata properties in the interburden and cannot be used to evaluate individual pillars. We apply the Rocscience RS2 two-dimensional finite element program as a complement to ACPS. Complex multiple-seam scenarios can be characterized by a modified pillar stability factor where the ACPS single-seam development pillar stability factor is multiplied by the ratio of average single-seam development stress to the stress in the critical pillar at the multiple-seam stage, using the values of o as predicted by RS2. Several series of test examples involving two-seam analyses with isolated remnants, gob-solid boundaries, and high extraction areas of varying width were used to calibrate simulated gob properties such that the ACPS+RS2 method generated results that matched a traditional ACPS Stability Factor to within 13%. As high-extraction areas become wider and deeper, progressively stiffer gob is simulated by increasing the Young's Modulus property in RS2. After achieving calibration with the test series, three examples of complex multiple-seam scenarios were evaluated, using reasonable properties for strata types and thicknesses based on core log data. Resulting modified stability factors are reasonably close to those predicted by ACPS, although affected by units of sandstone in the interburden and gob shadows superimposed on remnant structures. This tends to show that even in complex multiple-seam conditions, applying ACPS to the worst-case interaction should usually result in an acceptable pillar design.

Jun 25, 2024

By Robert B. Fulton

The objective of this chapter is to discuss the interrelationship between industrial minerals and chemical manufacturing. It is intended to supplement rather than duplicate the commodity chapters. Particular emphasis is given to the pertinent chemical element and to market factors. Condensing this broad subject into a few pages of this handbook permits treating only the most important elements derived from industrial minerals. Hydrocarbons, which quantitatively dominate as raw materials for the chemical industry, are omitted, as are the metallic elements and the minerals covered in other "use" chapters such as phosphorous, potassium, and nitrogen for fertilizers, and titanium dioxide for pigments. The remaining six elements of major importance are: boron, bromine, chlorine, fluorine, sodium, and sulfur. These elements are treated individually under separate headings. [Table 1] affords an overview of the main industrial minerals, the chemical products derived from them, and end uses of the products. Salt brines have particular importance as raw material sources for the chemical industry. Table 2 is a chart of the chemical compounds derived from four types of brines: (1) Owens Lake-type brines, which are sources of boron and sodium compounds; (2) Midland-type brines, from which bromine, iodine, and chlorides of calcium, magnesium, potassium, and sodium are derived; (3) Searles Lake-type brines, yielding boron, bromine, lithium, magnesium, potassium, and sodium compounds; and (4) Silver Peak- type brines, produced mainly for lithium. MARKET ATTRIBUTES Some of the important market traits common to industrial minerals used by the chemical industry are: 1. They are international commodities, such as fluorspar and sulfur, which largely move to foreign consumers. 2. Grade, and freedom from deleterious elements are important factors affecting their usability in chemical processes. An example is salt (NaCl) used in electrolysis where ultrapure evaporated salt is required to meet rigid specifications. 3. Purified products take on the characteristics of specialty items and command a distinctly higher price than the basic commodity from which they are derived. 4. In practically all cases, chemical users require some sort of cleaning or beneficiation of the naturally-occurring mineral to bring it to specification, and individual specifications may vary from user to user for essentially the same use. 5. In some instances it is necessary to strike a balance between what the vendor can supply and what the buyer requires, with the result that specifications have to be eased to afford the needed materials in marginal cases. 6. Because they tend to be bulk commodities, low cost for handling and transportation are important and such costs may limit the area from which a chemical user can draw his supply. 7. Shipments are usually in bulk and frequently in multiple-car, full-trainload or full-shipload lots to reduce transport costs, which in turn may require large terminal investment facilities. 8. Purchases are generally by contract of one year or longer term, with spot buying playing only a minor role. 9. Contract prices are usually fixed in short term commitments, but may vary according to assay, with premiums and penalties for content above or below the norm; however, general practice is for specifications to be fixed in the contract with minimums being set for the desired material and maximums for undesired elements. In longer term contracts, prices are often escalated on labor, fuel, and other vendor processing costs. 10. Suppliers of individual commodities to the chemical industry tend to be limited in number and are generally medium- to large-size producers that supply a few major consumers. 11. The bulk of the mineral volume is for basic chemical uses, sulfur suppliers to sulfuric acid producers and fluorspar for hydrofluoric acid producers being typical examples. These basic chemical products then are used for the production of other products. 12. Shortage of a supply of adequate quality leads consumers to seek substitutes. In the case of fluorspar, much work is being done on recovery of fluorine from phosphate rock. Success in the form of fluorosilicic acid and/or hydrofluoric acid production could, in time, affect the hydrofluoric acid chemical industry. 13. Markets tend to be characterized by cycles of shortage followed by oversupply, with attendant wide price fluctuations. 14. Baniers to trade can have an adverse effect on the necessary movement of industrial minerals used by the chemical industry in international trade. Antidumping laws, quotas, and tariffs can disrupt or dislocate normal markets. 15. Chemical industry consumers may back-integrate for security of supply or for favorable economics, sometimes by joint ownership and often with experienced mining partners.

Jan 1, 1994

By S. L. Bessinger

Introduction Extensive testing in the past ten years has shown that where a uniform natural gamma background is present in the strata bordering a seam, the thickness of the boundary coal left in place after mining can be determined by measuring the attenuation of that radiation (Nelson and Bessinger, 1989). Measurements made by the authors in underground mines in Pennsylvania, West Virginia, Ohio, Illinois and Kentucky and by others in Wyoming and New Mexico have shown the presence of such a gamma background (Nelson. 1989). Natural gamma coal-thickness sensors of several configurations have been tested in mines owned and operated by the Consolidation Coal Company (Consol) in Pennsylvania and West Virginia (Nelson and Bessinger, 1988). This paper describes the installation of a natural gamma coal-thickness sensor on an operating continuous miner. Previous tests had shown that the NGB-1000 coal-thickness sensor manufactured by American Mining Electronics, Inc., of Huntsville. AL, is an accurate, mine-worthy instrument. This large gamma detector consists of a sensing head and a control panel. The sensing head contains thallium-doped, sodium iodide scintillating crystal, which is coupled to a photomultiplier tube. The control panel contains the electronic components required for calibration, count conversion and display to the operator. Methods Conditions at a Consol mine in northern West Virginia require that 10 to 15 cm (4 to 6 in.) of coal be left at the roof boundary of continuous miner development sections. This roof coal is required because the shale of the immediate roof is friable and unstable. In the past, operators have used a dirt band that is usually visible near the top of the seam as a guide in maintaining the proper cutting horizon. However, this is not always reliable. Earlier observation showed that the actual thickness of the coal left on the roof varied widely; further, it was noted that occasional, accidental excursions into the immediate roof required supplementary roof control measures, such as installation of planks or center bolts. Thus, it was concluded that operators needed a better source of guidance for control of the cutting horizon, and a roof-coal thickness sensor was scheduled for installation. The NGB-1000 sensor was installed on a Joy 12CM10 continuous miner in June 1988. The sensing head was mounted on the cutter boom of the miner, and the control panel was mounted in the operator's cab. Power for the sensor was initially derived from an intrinsically safe battery power supply. Initial measurements with the sensor showed that the calibration was the same as that used in earlier tests at two other mines, indicating the uniformity of the natural gamma background above the Pittsburgh seam. Operating personnel were initially skeptical of the instrument's accuracy, and were hesitant to use its readings as a guide in maintaining a proper cutting horizon. Because gamma attenuation, the instrument's operating principle, is somewhat abstract, attempts to demonstrate the instrument's accuracy by explaining that principle were generally ineffective. It was found, however, that an operator could usually be convinced of the usefulness of the instrument by placing a large piece of coal of fairly uniform thickness over the instrument's sensing head and allowing the operator to see that the instrument reading increased by an amount very near his estimate of the thickness of the piece. The mine was provided with seven battery power supplies and a charging station. The charging station was kept in the lampman's office, and the mechanic on each shift was instructed that he was responsible for two battery power supplies each day: a freshly charged one to be taken in at the beginning of his shift and a depleted one to be brought out at the end. This system worked well for a few weeks, but eventually some battery power supplies were left in use so long that their batteries were discharged too deeply to allow recharging. In addition, transport and recharging of the batteries represented an additional task for the mechanics, who were already very busy. Consequently, a request was filed with MSHA to allow the sensor to be powered through intrinsic safety barriers by an electronic power supply connected to machine power. The permit was granted, and the sensor was connected to machine power. After the sensor was connected to machine power, the only operating problem experienced was occasional failure of cables. A supply of the required cables was made and delivered to the mine so damaged cables could be quickly replaced. Much of the cable damage could be eliminated by slight modifications to the miner during a rebuild, so that cables could be installed in more protected locations. After the sensor had been in operation for about two months, a survey was made to determine its effect on continuous miner operations. In previous research, coal thickness measurements made in 88 locations by the natural gamma method were compared to measurements made in the same locations by observing drill cuttings and by inspections of drill holes with a borescope. That research showed that the gamma method is at least as accurate as the other two methods (Nelson and Bessinger, 1989) and is also much easier to use. The object of the survey described here was not to assess the accuracy of the natural gamma measurements. but rather to determine the effectiveness of the sensor output as a guide for the operator in maintaining control of the cutting horizon. Thus a smaller, hand-held gamma detector

Jan 1, 1992

By John C. Ferm

Although roof falls account for approximately half of the fatalities in underground coal mining, the factors that produce them are not clear. Mining methods can account for some, but in others the controlling factors are the geological setting in which the mining takes place. The present investigation was undertaken to determine the kinds of rocks that were most likely to produce roof problems. Such knowledge would be helpful to mine planning since rock samples can be obtained in advance of mining by core hole drilling. The investigation was undertaken in underground mines in the Pocahontas #3 seam in southern West Virginia and southwestern Virginia. The Pocahontas #3 is one of the most widespread in this region and is a major source of high quality metallurgical coal. There are at least ten very large mines in the area and many smaller ones. The region has a very long mining history and considerable experience has been gained in roof control problems. The study was conducted with the support of the U. S. Bureau of Mines.

Jan 1, 1978

By W. S. Sadik

This paper presents the ventilation design for an underground testing facility in a salt medium. The underground workings consist of two-12 ft. (3.66 m) diameter watertight shafts and approximately 6000 ft. (1830 m) plus of drifting. These facilities are designed for the purpose of providing access for an in-situ testing of a salt horizon in which a variety of hazardous and non-hazardous materials could be temporarily stored or permanently deposited. The ventilation system includes a design basis for gassy mine conditions and provides for the option of bulk air cooling and heating.

Jan 1, 1988

By A. L. Wilder, S. N. Dixon

Introduction Coeur d'Alene Mines Corp.'s largest precious metals property is located in the historic Rochester Mining District 40 km (25 miles) northeast of Lovelock, NV. The property encountered cold weather operational problems soon after its fall start-up in 1986 due to its elevation of over 1830 m (6000 ft). The problem of ice buildup on the heaps because of sprayed solution application was faced immediately. It was felt that allowing ice to build up all winter long until a spring thaw was impractical due to the large area under leach. Further, the operating cost and delivery schedule for a solution heating system was unacceptable. The development and installation of a leach solution distribution system using drip emitters made efficient, cost-effective winter operation possible. Other benefits of this system have also been observed and are discussed here. General process description 15,422 kt/day (17,000 stpd) of - 1.27-cm (-1 /2-in) crushed ore from the three-stage crushing plant are delivered to the leach pad using 77.1 t (85 st) rear dump haul trucks. The ore is drifted into place with a D-9 bulldozer. Leach panels are contiguous and are approximately 8861 m'(90,000 square ft) in area built in 6-m (20-ft) lifts. New panels are built on top of older areas to a final height of 61 m (200 ft). Each panel is ripped and cross-ripped prior to leaching. Barren solution is distributed to the heap using drip emitters at rates of 0.02 to 0.41 L/min/m2 (0.0005 to 0.01 gpm per sq ft), depending on the age of the panels. The pH of the leach solution is 10.7 with a cyanide concentration of 0.75 kg/t (1.5 lb per st). Approximately 50% of the silver and 80% of the gold are finally recovered. Pregnant solution percolates though the heap and flows by gravity into one of two 9.46 ML (2.5 million gal) pregnant solution ponds. The solution is then pumped to a conventional Merrill-Crowe process plant. Clarification takes place in three 9464 L/min (2,500 gpm) capacity filters. The solution is then pumped to a packed vacuum deareation tower for the removal of dissolved oxygen. Typical deareated solution contains 0.7 parts per million dissolved oxygen. Precipitation of gold and silver is accomplished by adding a zinc dust slurry to the deareated solution at the suction of the filter press feed pump. Precipitated gold and silver are recovered in three recessed plate and frame filter presses. Barren solution is discharged into a 11.7 ML (3.1 million gal) pond where cyanide makeup occurs. This solution is pumped back to the heap for further leaching. The precipitate filter cake, containing approximately 75% dore (Ag + Au), is then fluxed with anhydrous borax, soda ash, sodium nitrate and fluorspar to yield a neutral, bisilicate slag. The fluxed precipitate is then charged into a propane-fired melting furnace and heated to 1150° C (2100° F) for 3 1/2 hours. Slag and dore bullion are poured into conical cast iron pots yielding buttons of 800 to 1000 troy oz. The dore typically contains 98.5% silver and 1 % gold. Slag is crushed and tabled to recover the trapped dore blebs and beads. Concentrate from the table is returned to the furnace. Table tails are sent to the crushing circuit and out to the leach pad. Solution application The area kept under leach at Rochester is approximately 130 000 m2 (1.4 million sq ft). Barren solution is delivered to the pad at 21.2 kL/min (5600 gpm) for a resultant application rate of 0.16 L/min/m2 (0.004 gpm per sq ft). A traditional solution sprinkling system using No. 12 Senninger Wobblers with individual pressure regulators was installed at the onset of leaching activities. The Wobblers were placed at 9.1-m (30¬ft) staggered centers and were fed off of a gridwork of Yellowmine plastic piping. Solution flow rates were moni¬tored to each panel. The onset of cold weather with an average nighttime temperature of -12° C (10° F) made it apparent that continual operation would not be possible with the sprinklers. A significant amount of ice was built up on top of the heap, making maintenance and pipe removal dangerous, if not impossible. Leach solution application was restricted to daylight hours to inhibit ice formation. Process plant flow rates were reduced to maintain steady-state operating conditions. However, as daylight temperatures dropped below freezing, ice continued to accumulate due to the sprays. Besides the obvious operating hazards brought on by the growing icefield, there was also the potential environmental hazard associated with an early thaw melting the ice too rapidly for the solution containment facilities. One other option for preventing ice formation was heating of the barren solution prior to spraying. Initial plant design allowed for expansion of the propane storage and distribution system as well as modification of the barren piping for a solution heater. This option was not exercised because the operating costs for an adequate system would have been prohibitive, and timely delivery of a system was not available. An investigation was conducted on the various drip irriga¬tion products available, since subsurface solution applicators would eliminate ice formation altogether. Systems utilizing external flow emitters were ruled out because of their ten¬dency to clog when buried. Emitter systems using perforated tubing were also eliminated from consideration due to their inability to adequately control flow over required lengths of tubing. An in-line emitter system was finally selected which demonstrated clog resistance and adequate flow control, enabling direct burial.

Jan 1, 1990

By Fred N. Kissell

INTRODUCTION During the last few years, several new ventilation developments have attracted the interest of mining engi¬neers. Some of these developments are applicable pri¬marily to hard-rock mining, while others are more applicable to coal mining. STOPPINGS Parachute Stopping The parachute stopping is a new type of quick-erect stopping that is intended for temporary use in hard-rock mines (Kissell, Thimons, and Vinson, 1975). As shown in Fig. 1, the stopping is shaped very much like an ordinary parachute, with a canopy of impermeable fabric that is sewn to regularly spaced straps running to a common point. To erect the stopping, the straps are attached to a fixed anchor point such as a roof bolt, and the edge of the canopy is lifted into the moving air¬stream. The airstream pops the parachute canopy into place, and the differential air pressure across the stop¬ping holds it in place, forcing the fabric against the walls, roof, and floor of the mine opening. The principal advantage of the parachute stopping is that it requires only a few minutes to install, making it a great time-saver for emergency use or for day-to¬day changes in ventilation during the production cycle. However, the parachute stopping does require some minimum air velocity to lift it and some minimum differential pressure to hold it in place. For a fabric weighing 0.27 kg/ m2 (8.0 oz per sq yd), the minimum air velocity is about 0.5 m/s (100 fpm), and the mini¬mum differential pressure is about 0.05 kPa [0.2 in. water gage (WG) ]. There is always some air leakage around the stop¬ping, mainly depending upon the extent to which pipes or other obstructions encumber the airway and prevent good sealing. Leakage of a few cubic meters per second (a few thousand cubic feet per minute) can be expected, unless foam is used to improve the seal at the edges of the canopy. Quick-Fix Blowout Stopping The quick-fix blowout stopping is a variation of the parachute stopping (Thimons and Kissell, 1976), and it is used in the proximity of blasting operations. This type of stopping is designed to be blown out easily by the blast forces, and it may be reinstalled quickly and easily. The long high-strength straps of the parachute stopping are replaced by groups of short straps that tear easily. These straps are attached at six equally spaced locations around the perimeter of the canopy. To erect the stopping, one strap of each of the six groups is fastened to the mine wall, roof, and floor by using spads, by setting pins with a powder-actuated gun, or by tying the straps to some firm anchor point. Once the straps have been attached, the differential air pressure across the stopping, which must be at least 0.025 kPa (0.1 in. WG), forces the stopping perimeter against the mine walls, thus creating the air seal. It is the self-sealing feature of this stopping that makes it a significant time-saver. Only a few attachment points are needed; in many cases, four attachment points are sufficient, since the stopping naturally tends to form a seal with the airway surfaces. When nearby produc¬tion blasting exerts excessive forces on the stopping, one or more of the straps tears away from its attachment point, protecting the stronger canopy from damage. Damage-Resistant Brattice The damage-resistant brattice is a stopping that is designed for use in mines such as salt and limestone mines where the differential pressures are low and the roof is relatively flat. As shown in Fig. 2, the damage-resistant brattice consists of a series of brattice panels that are hung vertically and joined by Velcro® connections. When the brattice is subjected to strong blast forces, the Velcro® connection peels apart and allows the panels to open without incurring damage. The Velcro® connections can be resealed by hand within a matter of minutes. Such damage-resistant brattices have withstood the blast effects of 318 kg (700 lb) of ammonium nitrate-fuel oil (ANFO) explosive detonated as close as 91 m (300 ft) from the brattice. Ordinary brattice cloth is used for the panels, with a 51-mm (2-in.) wide strip of Velcro® hooks sewn along one edge of the length, and a 51-mm (2-in.) wide strip of Velcro® pile sewn along the other edge. Both the hooks and the pile are sewn onto the same side of the brattice cloth. The resulting Velcro® seal formed be¬tween adjacent panels is perpendicular to the brattice itself, and the leading edge of the seal can be directed either toward or away from the blast forces; the brattice works equally well in either case. To hang the brattice, panels of brattice cloth about 0.9 m (3 ft) longer than the height of the airway are cut from a 1.8-m (6-ft) wide roll. The additional 0.9 m (3 ft) of brattice cloth allows 0.3 m (1 ft) for attachment to the roof by means of a board, with 0.6 m (2 ft) for forming a good air seal at the floor. Each brattice panel is wrapped once or twice around a 51 X 102 mm (2 X 4 in.) or 25 X 76 mm (1 X 3 in.) mounting board that is 254 to 305 mm (10 to 12 in.) shorter than the width of the panel. For convenience in

Jan 1, 1982

By R. Spring

Artificial intelligence, and especially expert systems, are currently receiving considerable attention. Of interest to the control engineer is the utilization of these tools for real-time process control applications. To evaluate this technology, a real-time expert control system was implemented in one of the grinding circuits at Brenda Mines Ltd. The performance of this micro computer based expert system compared satisfactorily to the proven supervisory control system currently in use. The study confirmed the expectations that expert systems have potential for real-time, supervisory control of mineral processing operations.

Jan 1, 1989

By Lok W. Home

HISTORICAL DEVELOPMENT At the beginning of the 1970 more than 90% of pro¬duction drilling in open-pit mining was obtained with the use of rotary drills. At the same time, the use of rotary drills underground was virtually nonexistent. Several progressive mines began questioning whether efficient rotary drilling could not be adapted to under¬ground applications. In the late 60s and early 70s at¬tempts to use rotary drilling were made by Inco, Ltd. and Meramac Mining, utilizing drill units designed as a combination of small diameter raise drills and rotary drills. These initial attempts were unsuccessful primarily due to the immobility of the equipment and low degree of dependability. At the same time, experiments con¬ducted by Inco and Meramac using large diameter in¬the-hole percussion drills produced immediate encourag¬ing results. The first serious attempt at utilizing rotary drilling underground was made by Texasgulf in 1971 at Kidd Creek mine in Canada. An experimental stope 15 m wide, 36 m long, and 30 m deep was drilled using 200-mm diam rotary drilled holes. The machine used was a Robbins' Model 11 D. The Model 11D is a rotary drill designed to drill drain and utility holes under¬ground. The results of this test revealed that large diameter blasthole stoping was feasible. The indications were that the major areas of concern-fragmentation and pillar damage-could be controlled and a new low-cost stoping method resulted. At this point Texasgulf and Robbins entered into a cost-sharing agreement to develop a rotary drilling unit specifically for this stoping method. Drilling of the trial stope required pouring a concrete pad at each drill hole plus anchoring the derrick to the pad by means of rock¬bolts. Utilization expressed in terms of drilling time to available time was less than 20%. It was obvious that the required unit had to be mobile, setup time had to be minimized, and rod handling had to be improved. The first attempt at an underground production-type rotary drill consisted of mounting the basic 11D on an air-powered Eimco 600 Series crawler. The crawler was equipped with four screw-type leveling jacks operated by an impact wrench; rod rack facilities for 60 m of 1.2-m long, 170-mm diam rods; a rod handling facility consisting of an air-lift swing boom and a pipe loader which automatically positioned the rod under the float¬ing drive box; a turnbuckle anchor system requiring in¬stallation of four rockbolts in the drill drift floor and placing turnbuckles from the rockbolts to the corners of the drill; and trailing power units containing the electrics and hydraulics required to operate the drilling unit. Operation of this unit revealed a vast improvement in drilling time but some problem areas remained. The anchoring system required secondary drilling and proved difficult to adapt to the localized irregular floor surface and the trailing power unit was difficult to maneuver and was causing unnecessary maintenance costs. Addi¬tional improvements resulted in the standard 11MD rotary blasthole drill. DESCRIPTION OF UNIT This standard model consists of the drill and the power unit integrally mounted on an air-powered crawler chassis (Fig. 1). The rod-handling and leveling¬jack systems are similar to those used on the prototype Texasgulf unit and remained relatively the same. The anchoring system has been replaced by the roof jack, initially screw type but now hydraulically actuated. The anchoring system is necessary as the drill exerts more thrust than the weight of the unit. During operation it is found that the two front jacks adequately anchor the machine. The complete unit's operational dimensions are: height, 3.2 m; width, 2.1 m; and length, 4.2 m requiring a 3.5- x 3-m drill drift. The derrick is pivot mounted and raised and lowered by a hydraulic cylinder reducing the transport height to 2.5 m. The derrick is also turnbuckle mounted to allow dip angle drilling from vertical to 0.52 rad (30°) off vertical. The drill derrick can be mounted to allow dip axis to be parallel or perpendicu¬lar to the long axis of the unit. The standard rods are 1.2 m in length and are contained in racks of 10 which can be removed for transportation. Drilling torque is produced by a 56-kW (75-hp) electric motor driving a double vane pump which in turn drives a Robbins high torque gear-type hydraulic mo¬tor. In addition to the hydraulic motor, the machine drive system has a 3.9:1 planetary gear reducer. This results in six speeds with the following minimum and maximum range: Maximum rpm 112 Torque 3725.5 J (380 kg m) Minimum rpm 10 Torque 14 706.0 J (1500 kg m) Drilling thrust is provided by two bootstrap cylin¬ders pressurized by a variable volume pump. This sys¬tem can produce from 0 to 267 231.2 N thrust in either direction. Bailing air (approximately 17 m3/min at 382.2 to 490.3 kPa) is introduced through a rotary swivel pass¬ing directly through the center of the drive system to keep efficiency and losses to a minimum. Hole annulus is also kept to a minimum by using 170-mm rod when drilling a 200-mm hole. A small amount of water is added to the air for dust suppression. The machine has to be supplied with air, water, and electrical power. It does not appear feasible to provide an integral air power supply with the rotary drill such as is done with open-pit rotary drills. Air is required and installed during the primary development of the drill drift. The machine can be adapted to drill hole diameters from 170 to 228 mm for blasthole operations. The drill string consists of the rotary bit, crossover stabilizer, roller stabilizers, and standard rods. The operator controls speed and thrust pressure from an operating console which can be mounted in a conve¬nient position. Mount Isa has made an adaption to the operating console to permit remote collaring of drill holes in an uncleared floor.

Jan 1, 1982

By J. R. Allen

Environmental laws and regulations have proliferated over the last two decades. 1/ Within the last year, Congress has dramatically expanded the scope of the Superfund program 2/ and has enacted significant new amendments to the Clean Water Act.3/ In addition, EPA has shown no signs of slackening the pace at which it promulgates new regulations to implement statutory programs. The agency projects that over the next few years it will issue approximately 275 new regulations under the statutes it implements. 4/

Jan 1, 1988

By J. Lorosch

Improvement of current leaching technology was found necessary when processing ores which are more difficult to leach because they generate highly viscous, air-rejective pulps. Cyanidation of these pulps has to run at low oxygen concentrations if the conventional aeration technique with compressed air is used. Since the dissolution rate of precious metals mainly depends on the concentration of dissolved oxygen in the pulp [1, 4, 5], the kinetics of the gold extraction become slow in such cases, which in turn often leads to a significant decrease in the overall gold output within the conventional leaching time of up to 48 hours.

Jan 1, 1988

By R. E. Greuer

The design of mine fire emergency plans requires that the contamination of mines by fumes, and the mutual interaction of fires and ventilation systems be pre-calculated. Several computer models were developed by the Bureau of Mines for this purpose during the last decade. The models advanced in sophistication from the transient-state simulation of fume concentrations for the early stages of a fire to the steady-state simulation of airflow rates, concentrations, and temperatures for fully developed fires, and finally to the complete transient-state simulation of fires and ventilation systems. The principal features of these models are described.

Jan 1, 1989