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By Murray Malott

The Sheppard Subway extension is currently being constructed in Toronto, Canada for the Toronto Transit Commission (TTC). McNally International Inc., in a joint venture with PCL and the Foundation Company of Canada were awarded the Twin Tunnels contract of the Sheppard extension. The Twin Tunnels project is approximately 4.4 km in length and consists of two separate twin bore drives each 3.3 km and 650 m in length and approximately 400 m of cut and cover construction. This paper presents a case history of the construction of the soft ground tunnels through the complex Toronto glacial deposits using two 5.9 m outside diameter Lovat earth pressure balance (EPB) tunnel boring machines.

Jan 1, 1999

By R. J. Smith

Tar sand deposits in Utah represent more than 25 billion barrels of in-place bitumen, 96% of the known U.S. reserves. The technological development of a hot water processing strategy for the Canadian tar sands and the commercial Production of synthetic crude oil from this natural resource suggest that similar success may be possible for the utilization of Utah tar sands to provide our country with a new domestic source of synthetic crude oil. In this regard an active research program to develop a hot water processing strategy for Utah tar sands has been sustained at the University of Utah for the past five years. The hot water process for Utah tar sands differs significantly from that used for Canadian, tar sands due to the inherent differences in the respective bitumen viscosities and the nature of the bitumen-sand association. These differences have led to the identification of certain design criteria necessary to achieve satisfactory phase disengagement during digestion. (8,9) Subsequent flotation of the digested tar sand is a subject of current attention and is considered in this contribution for the Asphalt Ridge tar sand deposit. Although contact angle measurements of pure Asphalt Ridge bitumen indicate moderate hydrophobicity, air bubble attachment to the bitumen concentrate taken from a 37 liter flotation cell is not possible. This surprising result suggests that the flotation separation is dependent on air bubble entrapment rather than on attachment due to surface hydrophobicity. The occlusion of air bubbles in the bitumen is apparent from visual examination of the concentrate, especially at lower temperatures. It seems that a bitumen bubble agglomerate forms in the impeller region of the flotation cell, the effective density of which is such to allow the agglomerate to float to the surface. A factorial design of the major operating variables in the flotation separation indicated that the quality of separation was significantly dependent on the flotation temperature and to a lesser extent on the degree of agitation. For a flotation temperature of 77C recovery of 96.7 percent was realized at a grade of 61.0 percent bitumen. The improved separation at higher flotation temperatures was found to be due to the decrease in bitumen viscosity resulting in more effective rejection of coarse sand from the concentrate.

Jan 1, 1980

By Betty L. Gibbs

Microcomputers are becoming more convenient to use at all levels of the business world, including the mining industry and mine offices. The state of the mining industry requires that we become more productive and innovative in our analysis of deposits and operations. With the help of computers, examination of options to increase productivity can be performed quickly and inexpensively. Effective use of the computer for designing underground workings combines the calculation power and speed of a computer with the technical expertise and judgment of the engineer. Computer programs can perform calculations in minutes that could take an engineer several months. The engineer must be on guard when using the computer for any calculation because it is very easy to assume the results produced by a computer program are always correct. The computer is a sophisticated calculator but cannot recognize logical errors nor transposition of numbers. Also some intuitive or obvious decisions made by an engineer are difficult or impossible to program effectively.

Jan 1, 1987

By Gordon C. ‘Bud’ Presley

The pumice and pumicite review arrived too late for the June issue. Estimated domestic production of pumice and pumicite in 1992 was up compared with 1991, according to US Bureau of Mines statistics. The Bureau estimate of domestic production of 400 kt (440,000 st) for 1991 turned out to be 401 kt (442,000 st) of actual production, the lowest since 1989. In 1992, the estimated the value of pumice and pumicite sold or used in the United States was $11.6 million, up only $200,000 from the 1991 value. There was an increase in the mine, or mill, average price from the five-year low of $11.44/t ($10.37/st) in 1986 to $26.42/t ($23.96/st) in 1992. Estimated tonnage produced in 1992 was 439 kt (484,000 st). Apparent domestic consumption was 642 kt (707,000 st), up 135 kt (149,000 st) from 1991. Imports for consumption were estimated to be 215 kt (237,000 st) last year. Domestic production supplied 68% of the demand in 1992, with net imports supplying the remaining 32%. This compares with domestic production sup¬plying 53% and net imports supplying 47% of the apparent consumption in 1988. Exports were about 900 t/a (1000 stpy) for the period 1984 to 1988. In 1989, exports rose sevenfold to 7 kt (7700 st) and to 20 kt (22,000 st) in 1990. The Bureau of Mines reported that the amount of pumice and pumicite exported in 1992 was 12 kt (13,200 st). It is assumed that these exports were from the Pacific Coast because imports, mainly from Greece, continued to be a market factor on the East Coast and the Gulf Coast States in the United States.

Jan 1, 1993

By E. O. Mixon, Jack D. Kennedy

INTRODUCTION This contract is the most northerly portion of the Metropolitan Sanitary District of Greater Chicago's Tunnel and Reservoir Plan. (TARP) The project begins at Addison Street in Chicago and proceeds northerly approximately ten miles to end at Wilmette Harbor and Lake Michigan. The project which consists of 2 work shafts, 32 drop shafts, 7.322 km (24,020 ft.) of 6.7 m (22 ft.) diameter circular rock tunnel, 8.458 km (27,750 ft.) of 9.2 m (30 ft.) diameter circular rock tunnel and miscellaneous rock bolting and grouting, was bid for a total cost of $63,140,480. The tunnel invert elevation varies from 65.5 m (215 ft.) to 77.4 m (254 ft.) below the ground surface, which places the tunnel in a rock formation of dolomite limestone. At the tunnel invert, the unconfined compressive strength of the limestone varies from 96 MPa (14,000 psi) to 152 MPa (22,000 psi). The 6.7 (22 ft.) diameter tunnel required the mining of 10.9 cu m. (14.2 cu yd.) of rock per lineal foot of tunnel while the 9.2 m (30 ft.) diameter tunnel required the mining of 20.1 cu m (26.3 cu yd.) of rock per lineal foot of tunnel. The total cubic meters of rock required to be excavated from both tunnels was 818,821 cu m (1,070,978 cu yd.). All of this will be mined by tunnel boring machines (TBM). These tunnels were not to be lined with concrete. The thirty two drop shafts have upper portions that are cylindrical in the overburden and rock sections. They were divided into two groups, depending on diameter, with a different construction method for each group. The first group consisted of twenty three of the

Jan 1, 1979

By Nick Barton, Khosrow Bakhtar

Two plate load tests of 1 m2 area were per- formed in a pressure tunnel at the Rocky Mountain Pumped Hydro Project. Reaction for the 250 tons loading in this 10.5 meter diameter tunnel was provided by deeply anchored, diverging rock bolts, eight per load frame. Pressure build up to 2.5 MPa was provided by circular flatjacks. Rock deformation was recorded at depths of approximately 0.3, 0.9, 1.2, 1.8 and 3.3 meters using axially located multiple position borehole extensometers. The tests were performed in representative units of massive shale and thin to medium bedded siltstone. Stress was built up in four equal steps to 2.5 MPa, held constant for one week for creep monitoring, then unloaded, reloaded and unloaded over a period of one day. The most deformable unit was the massive shale which exhibited 1.3 mm of permanent deformation at the surface after the first loading and unloading cycle, which included 0.36 mm of creep. After the second loading-unloading cycle, the permanent deformation had increased to 2.6 mm, following a maximum deformation at the surface of approximately 3.6 mm. At the maximum stress of 2.5 MPa, the tangent modulus averaged 0.4 GPa when loading, and 0.3 GPa when unloading. The thin to medium bedded siltstone exhibited only 0.17 mm of permanent deformation at the surface after the first loading-unloading cycle, which included only 0.05 mm of creep. After the second loading-unloading cycle, the permanent deformation had increased to 0.27 mm, following a maximum deformation at the surface of approximately 0.53 mm. At the maximum stress of 2.5 MPa, the tangent modulus averaged 6.2 GPa when loading and 21.9 GPa when unloading.

Jan 1, 1986

By J. D. Miller

The sedimentary western phosphate deposits of Idaho, Montana, Utah, and Wyoming tend to be calcareous and generally carry a significant amount of dolomite along with collophanite. In some cases, such as the Smokey Canyon Mine, Afton, Wyoming, the ore will also contain mudstone, and in this case the aluminosilicate minerals (clay, feldspar, etc.) together with some dolomite are floated from the collophanite in column flotation cells with an amine in order to meet product specifications (31 % P205) (Service et al.). In other cases, such as the SF Phosphates Limited Mine, Vernal, Utah, the ore is treated by traditional fatty acid flotation of the collophanite with mechanical flotation cells (Weber, 1986). Preliminary tests have now been conducted to evaluate air-sparged hydrocyclone (ASH) technology for the flotation of western phosphate ore. For this preliminary study, samples were taken from the hydrosizer overflow stream (18-21 % P205) at the SF Phosphates Limited Plant in Vernal, Utah for scavenger flotation testing with the 2-inch ASH system in laboratories at the University of Utah. Selected design and operating variables are being studied, and the initial results show that satisfactory separations can be achieved by high-capacity ASH flotation at appropriate operating conditions. For example, with single-stage ASH flotation 80% recovery can be obtained with a scavenger concentrate grade of about 26% P2O5. These results are comparable to, or better than, those obtained at SF plant operations with conventional flotation cells and suggest that ASH flotation technology can be used to good advantage in the phosphate industry. Of course, research should be continued to consider other process streams, optimize the separation efficiency, and evaluate larger diameter ASH units.

Jan 1, 1993

By R. C. Richardson

Barite Worldwide barite production for 1985 decreased 1.3%, while US barite production increased 1.4%, according to US Bureau of Mines statistics. The increase in US production is attributed to heavy demand for oil and gas drilling during the first quarter and an unrealized heavy expected demand in oil and gas markets for the last quarter. Nevada accounted for 82.1% of the 807 kt (890,000 st) of barite produced in the US last year, according to the Bureau. Other producing states were California, Georgia, Illinois, Missouri, Montana, Tennessee, and Washington. In 1985, the US imported 1.8 Mt (2 million st) of crude barite and 64 kt (71,000 st) of ground barite. China led the way with 43.3% of the total. Imports from India accounted for 21.6%; Morocco, 15.5%; and Thailand, 5.6%. The 1.8 Mt (2 million st) of crude barite imported is an 18.8% increase over 1984. Apparent consumption of barite for 1985 was about 2.6 Mt (2.9 million st), a 15.7% increase from 1984. Due to very low drilling rig activity during the last quarter of 1985, actual consumption is probably less as crude barite inventories at mines and plants increased. According to the Bureau, 90% of barite consumed in 1985 was for the oil and gas market. Paints, rubber, glass, and other chemical uses consumed the other 10%. With the price of oil falling dramatically early in 1986, the outlook for oil and gas drilling looks to be significantly lower than 1985. This lower demand for barite, coupled with the high year-end inventories, should lower both imports and domestic production. Significant lower prices on imported barite are expected, especially from China, Morocco, and India. This also could further lower US production of barite this year.

Jan 5, 1986

By Agne Rustan

INTRODUCTION This work is a continuation of the preparatory study by Rustan and Stöckel 1979 (3) and (4) of the possibilities to introduce automatic mine ventilation in Swedish mines and construction works. The preparatory study has been financed by the Swedish Board of Technical Development. In the study it was recommended that an automatic step by step control of the fans near the working place due to diesel exhaust emission would be the best solution. A more advanced alternative would be to regulate the fresh air continuously in proportion to the contamination. The latter alternative has been studied by Boliden Mineral Co in the Udden mine. In this study it was found that the air supply could be regulated continuously but it was very important where the sensors feeling the contamination were placed. An intense study of the environment underground is now being planned by our division together with specialists from the National Board of Health and Safety, The Swedish Work Environment Fund, LKAB and the local Governmental Inspection of the Workers Safety. Different gases, radon, dust, airspeed, moisture temperature and pressure is planned to be measured continuously and correlation between the parameters is going to be established to decide the best measurement parameters for regulating the system. The best situation of sensors is also going to be studied. Despite this work hasn't been done yet this paper covers some of the problems which could be encounted with continuous regulation. In this report the environment has been studied during the test with step by step regulation of fans by radio. Diesel vehicles are the largest polluters under ground in Swedish mines and we have therefore suggested that the large diesel vehicles should regulate the fans. The goal of this study has therefore been to find how active diesel vehicles should be able to start and stop gallery fans at both intake and return air. In this paper different methods for activating fans are discussed. Then the best methods are chosen and tested in the mine. In this study we have found that changes of temperature is a cheap and effective method to start and stop cross cut fans in sublevel caving and radio signals seem to be an attractive system for start and stop of gallery fans. This work has been done in close collaboration between LKAB in Kiruna and the University of Luleå - Division of Electronics and Division of Mining and Rock Excavation. A reference group for the project has been established with ventilation experts from the mines, representatives for the local mine worker's unions and authorities from the local Governmental Inspection of the Workers Safety. VENTILATION SYSTEM STUDIED In the project the ventilation system at LKAB in Kiruna has been studied. The mining method used is sublevel caving with 9-13 m slice height. The air is heated winter time over ground by hot water heat exchanger. The main fans are pushing air down to a main distribution level along the 4000 m long and about 100 m wide orebody. From the distribution level, air is taken through shafts down to the sublevels, where it is pushed through the levels by local gallery fans on both intake and return side, see fig 1. The air ( ~ 8 m3/s) is moving freely through the gallery head. The cross cuts are ventilated with smaller fans, diameter 500 mm, and flexible duct, see fig 2. The aim of this study has been to study how to start and stop gallery- and cross fans and examine the environment when the fans are stopped . In Sweden there are no regulations telling that a minimum of air amount must be supplied. Instead the hygienic limits must be fullfilled.

Jan 1, 1981

By V. B. Menon, L. D. Michaels, M. E. Mullins

Introduction The cleaning of coal to remove sulfur and inorganic impurities has, in recent years, become an essential operation in any coal utilization process. US coals, on average, contain about 3% total sulfur, of which more than 60% is generally inorganic or pyritic sulfur. Existing technologies such as heavy media cycloning, selective agglomeration, and electrostatic separation have the potential for removing about 80% of the pyrite from coal (Khoury, 1981). In view of the large volumes of coal being used today, there exists a need for the development of new processes to further improve the separation of pyrite from coal. Colloidal particles in suspension possess surface charges that are related to the chemical composition of the particles. In aqueous suspensions, the surface charge is a function of pH. For coal-pyrite mixtures, it is possible, by an adjustment of the pH (also a function of polar molecules), to reach a state where the charge on coal is very close to zero, but the charge on pyrite is still substantial. The introduction of collector particles with a charge opposite that of the pyrite should then result in the selective flocculation of pyrite and collec¬tor. These flocculates should settle out much faster than coal and could be removed from the bottom of the settling column. This note discusses our preliminary observations of the effect of clay-based collector particles, especially bentonite, on the settling behavior of coal and pyrite. These experiments were conducted with a view to developing a new electrokinetic flocculation process for separation of coal-pyrite mixtures. Experimental The objective of the experiments was to determine the effects of clay-based additives on the settling rates of clean coal, pyrite, and coal-pyrite mixtures. Illinois Colchester (No. 2) coal was obtained after processing in an existing coal cleaning facility. This coal was reported to contain less than 0.3% pyritic sulfur, which was in the form of a fine powder with an average particle size of 5µm. Pyrite (of average particle size 5 µm) was also obtained from the same facility and reportedly contained less than 0.1% coal. The zeta potentials of coal and pyrite in water were measured as a function of pH using a zetameter. The pH was adjusted by adding sodium hydroxide or hydrochloric acid to the suspensions. Sedimentation experiments were conducted in 50-mL glass cylinders, and settling rates were measured by noting the height of clear liquid above the settling front as a function of time. Settling behavior was also charted by periodically taking photographs of the system. The following systems were investigated: •a slurry consisting of 2.5 g of clean Illinois Colchester coal in 50 mL of water of pH 6.0, with and without the addition of 0.5 g of bentonite (Fisher, purified grade) ; • a slurry consisting of 2.5 g of pyrite in 50 mL water of pH 6.0, with and without the addition of 0.5 g of bentonite ; and • a slurry consisting of a mixture of 1.25 g clean coal and 1.25 g pyrite in 50 mL water of pH 6.0, with and without the addition of 0.5 g of bentonite. In addition, experiments were conducted with kaolinite (Bath, South Carolina) and alumina (Fisher, purified grade) as collector materials to verify the observed effects. Results and discussion Figure 1 shows the zeta potential values for the Illinois coal and pyrite as a function of pH. The zeta potentials of coal and pyrite decrease with increasing pH, changing from an initially positive value at low pH to a negative value at high pH. The point of zero charge (PZC) for the coal is at a pH of 6.2, while the PZC for pyrite occurs at a pH of 10.3.

Jan 1, 1988

By G. Sanders

Discussion by G. Sanders The studies reported on in this paper were initiated to draw attention to the severe contamination problem in the Section 30 drilling program at Chimney Creek. The lithologic-subset sampling study reached a different conclusion from that presented in your paper, and I wish to comment on your subsequent analysis of the data and your conclusions. Request for more complete data In the section on subsampling, you mention that the subordinate lithologies were separated and sampled, yet only the dominantlithology gold value is plotted in Fig. 4. In a contamination study, the reader is interested in the assay values for the individual subsets. Please include a table of the subsample assay data in your reply. Also, please indicate which analytical methods were used to arrive at the gold values in the subsampling study. Turning barren rock into low-grade ore Figure 5 is very revealing and typical of all of the cross sections in Section 30. Note the long strings of low-grade mineralization spread out for hundreds of feet below the ore zones. There were some very high gold values found in certain contaminated fractions during the subset sampling. The conclusion, here, was that the distinctive, strongly-mineralized dolomite layer was probably loose and crumbly and continued to disintegrate during drilling. This caused salting of the unmineralized rock samples below. Missing the high-grade part of the ore body In your statistical analysis, you directly compare the reverse circulation assays to the diamond drill assays in Section 30. Two points argue against a direct comparison and suggest the differences are greater than the 3 % that you report. First, any core loss in a gold zone most likely means that the true gold values are greater. The drillers lost significant amounts of the clay-rich, Section-30 gold mineralization. Also, the initiated salt-mud system, an attempt to improve the core recovery, met with little success. Second, the practice of not sampling core geologically, but instead sampling on even 5-foot intervals, adds a deliberate dilution to the core assay values by including a portion of nonmineralized rock in the first and last samples of each high- grade intercept. The result is often a pair of low-grade assay values on either side of a high-grade gold zone. In reality, a high-grade gold zone has a very sharp assay wall that is often bounded by barren rock. This sampling method may make the diamond-drill core assays more like the reverse circulation values and may help explain the statistical similarities you found. However, it does not represent the true gold values in the high-grade parts of the deposit. You cannot deny that, by careful geological sampling of the drill core, higher and sharper assay values will be obtained. The low core recovery and the diamond-drill-core sampling method used act together to lower the diamond-core assay values. The 3% difference you found between the reverse-circulation and diamond-core assay values could be much larger when you consider what the true diamond-drill core values would be with optimum core recovery and a geologic sampling method for the core. Should statistics have been applied here? The statement "... that reverse circulation holes have overestimated the values of some ore zones and underestimated the values of others" (p. 345) is not correct. The subsampling confirmed what the cross sections hinted at in Section 30. Namely, beneath the high-grade zones, the reverse circulation holes created, by contamination, large intercepts of low-grade ore in regions of barren rock. Because the low-grade material was not there to begin with, this is not a process of overestimating low-grade mineralization. The next statement that "the average result is similar to that of the diamond drill holes" may apply to the data set numerically, but it is not true when viewed spatially on cross sections. Adjacent reverse circulation and diamond drill holes are almost impossible to correlate, high-grade zone values vary widely and many low-grade intercepts make no geologic sense. The subset sampling and cross sections presented in the first part of the paper show that the reverse circulation portion of the data set has some serious problems, as highlighted above, and should not have been dealt with statistically at all. Conclusion Each ore body is different, and each drilling method presents unique sampling problems. In this case, the diamond drill is the

Jan 1, 1994

By A. E. Foscolos

British Petroleum (BP, 2005), indicates that world reserves for oil to be 1198 billion barrels, for natural gas the equivalent of 1156 billion barrels of oil and for coal the equivalent of 3219 billion barrels of oil. If we take into consideration the mean aver-age annual increase in demand of the last three years for oil, 2.6%, for natural gas, 3.9%, and for coal, 6.6%, then their respective reserves? lifespan are calculated to be 27, 34 and 38 years. Without taking into account their annual increase in demand, lifespan increases to 38, 67 and 161 years respectively. Since zero increase or perpetual increase in energy demands on finite resources is unrealistic the selection of the most suitable energy commodity for power generation rests upon the law of supply and demand of each commodity. This defines its price hence its selection for the production of cheap electricity. When the ratio of supply to demand is less than one, peak production is achieved. Any further increase in demand results in price increase to quench its demand. The Association for Oil and Gas Peak (ASPO) has calculated that when oil demand exceeds 84 million barrels/day, Peak Oil Production will be attained thus prices will start to increase in order to even demand with supply. Correspondingly, Gas Peak Production is expected to take place when demand exceeds 9.1 or 10.2 billion cubic meters/day. Peak Oil production has been calculated to take place in the current year, 2006, while Gas Peak production is estimated to take place between 2020 and 2025 provided that gas will not cover the energy deficit created from the lack of oil otherwise Gas Peak production could arrive as early as 2010. The use of coal for power generation is considered a better alternative than oil and gas if new combustion technologies with an energy efficiency of 43%, Germany, and 47%, Denmark, are adopted. These not only extend the lifespan of coal reserves but reduce also CO2 emissions by 25% and 31%, respectively. In the very near future CO2 sequestration and fixation in deep geological strata will eliminate its emissions thus rendering the use of coal the cheapest feedstock for power generation.

Jan 1, 2006

By John L. Halls

INTRODUCTION In this section, it is assumed that the reader is an engineer with sufficient training in underground mining work, preferably with budgetary control, to understand the mining methods outlined in other sections of this handbook. The mining-method descriptions are not repeated herein; only sufficient information is provided to enable the reader to comprehend the development of the costs presented. All costs are in 1977 US dollars, except where explicitly stated otherwise. CAPITAL INVESTMENT The capital investment in the underground mining portion of any mining property includes the following: 1) The development of original access involving adits, shafts, declines, etc. It includes the necessary equipment for transporting men, materials, and rock. 2) The purchase of mine equipment including loaders, underground haulage equipment, fans, pumps, electrical distribution, surface facilities associated with the underground mine, etc. 3) Sufficient preproduction mine development to assure steadily increasing production after startup. De- pending upon the mining method used, the preproduction development should provide a supply of three months to one year of ore that is developed and is ready to be stoped. Underground development also includes excavations other than shaft stations, loading pockets, sumps, etc., which usually are associated with the means of access and are excavated by the contractor. These excavations could include a crusher cavern, an ore- storage pocket, and underground workshops that may or may not be included by the shaft-sinking contractor. Portions of these could be delayed until after production begins, but it is preferable to complete them during the development phase, so there is no interference with the production buildup. 4) The staff and labor costs for all phases from inception through final production. 5) The provision of mine-surface equipment and facilities strictly applicable to the underground portion of the operation. The required facilities include a changehouse with a lamp room, a first-aid station, compressors, an electrical substation, workshops, a ware- house, water pipelines to the concentrator, water storage tanks, explosive storage, etc. A concrete-batch plant, a sand-fill plant, and boilers for heating may be required. 6) Other direct and indirect costs. 7) Engineering, design, and procurement costs. 8) Contingency allocations. 9) Contractor's fees, startup costs, and owner's costs. Original Access As part of a study for the US Bureau of Mines, Dravo Corp. conducted an extensive literature research survey covering underground operations in the United States between 1963 and 1973. Table 1 is compiled from the section of that study pertaining to the types of access to underground mines. Table 1 does not include distances attributable to the deepening of existing shafts, nor does it include shafts bored by raise drills or downhole machines that need existing underground workings for the disposal of the cuttings. During the period under review in the report ( 1963 through 1973). concrete-lined circular vertical shafts were the predominant means of providing the first access to a new underground mine; these shafts accounted for 80% of the original access in terms of the number of accesses completed, and 75% in terms of the total verti- cal linear distances. The inside diameters of the conventional circular shafts ranged from 3.7 to 7.0 m (12 to 23 ft). Most (72% ) of these conventional shafts were for production,

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 J. Drelick

Among the wide variety of modern surface chemistry analytical techniques, contact angle measurements still remain attractive for wetting characterization of solid surfaces. The captive-bubble and sessile-drop techniques are the most popular methods for the examination of flotation systems. Systematic measurements of advancing and receding contact angles for several three-phase systems using the captive-bubble and sessile-drop techniques show that the contact angle measurements depend on the technique used and may differ significantly for some heterogeneous and rough solid surfaces. Importantly, it was found that these discrepancies in contact angle measurements depend on the bubble (drop) size. Contact angles as measured with the static captive-bubble technique are compared with those measured with the dynamic captive-bubble technique, in which the capture of a bubble at the solid surface results after buoyant transport and attachment The dynamic captive-bubble technique allows for the formation of much smaller bubbles than is possible with the static placement technique in which a bubble is placed at the solid surface with a microsyringe. Although this dynamic technique cannot be recommended for the systematic wetting characterization of solid surfaces, its significant advantage is the ability to examine the contact angle/bubble size relationship over a wide range of bubble sizes, from several micrometers to several millimeters in diameter.

Jan 1, 1995

By Vittorio Manassero, Antonello de Risi, Filippo Cavuoto

"This paper deals with an innovative mechanized method to bore and simultaneously line circular shafts, adopted in the Naples Metro project for the installation of 10 vertical shafts for ventilation purposes. With this method, excavation of shafts is carried out under a positive head of stabilizing fluid by a milling machine, temporarily set at the bottom of a permanent lining made of precast reinforced concrete segment rings. The debris is evacuated by inverse fluid circulation. The shafts in Naples are 4.5 m I.D., from 34 to 44 m deep and they are all located downtown in a highly urbanized area. They were bored and installed into volcanic soils under a hydrostatic head of up to 30 m. Displacements and groundwater level were constantly monitored to keep control of the effects on the surrounding buildings and utilities. INTRODUCTIONThe extension of Metro Line 1 in Naples (Italy) from the previous terminal of Piazza Dante to Piazza Garibaldi is also called the “low stretch”. Most of it runs along the coastline, encompassing the historical city center, the Greek-Roman Neapolis. The Metro Line crosses under densely urbanized areas for the whole way (Figure 1). It runs beneath Via Toledo, which follows the layout of the ancient Aragonese walls. The first station (Toledo) is located halfway along this street. Then, after a wide bend, the line reaches Piazza Municipio, where the second station is located. From there, the line moves along the axis of Via Depretis and Corso Umberto I, where the stations Università and Duomo are situated. The fifth station is in Piazza Garibaldi and the end of the stretch is located at the Naples Business Center. The twin single track tunnels were bored into volcanic tuff by TBM. This design choice simplified by far the construction of the tunnels and reduced to practically zero the interactions between tunnelling and surface buildings and utilities. Nevertheless, this choice entailed that the line had to be located at depths to the order of 35-45m below ground level and 30-35m below the water table. At each tunnel segment between stations, a pair of ventilation chambers were constructed (10 in total). Each ventilation chamber is connected to one TBM tunnel by a vertical shaft and a horizontal drift at the base of the shaft. The shafts, in addition to tunnel ventilation, provide a safety access/escape."

Jan 1, 2016

By James R. Arnold, Cindy S. Jones, Michael F. Gleason, John O. Marsden, John G. Mansanti

INTRODUCTION The Chimney Creek Gold Mine (Gold Fields Operating Co. - Chimney Creek) is located 47 miles northeast of Winnemucca, Nevada, at the northern end of the Osgood Mountains. The operation is a wholly owned subsidiary of Gold Fields Mining Corporation, the North American branch of Consolidated Gold Fields PLC, London, England. The plant started up in November, 1987, less than three years after discovery of the orebody and three months ahead of schedule. Ore is mined in an open pit and is processed by combined dump leaching and milling techniques for gold and silver recovery. The mine is set to produce approximately 150.000 ounces of gold and 50,000 ounces of silver per year over a 12 year life at current reserve estimations. The mine was designed and constructed at a cost of $79.3 million with engineering and construction services provided by Davy McKee Corporation, San Ramon, California. Key Gold Fields operating staff were involved in the design of the facility from the start of the project: The Mine Manager, Plant Superintendent, Plant General Foreman, Maintenance General Foreman and Chief Metallurgist were all involved full time on the project within 5 months of the first ore discovery. Emphasis was directed at optimizing operating efficiency and in particular minimizing labor costs in the plant. It was recognized that a high level of instrumentation and control would be required to achieve this. The risk associated with the instrumentation and control systems implemented was to be minimized by using equipment and systems that had been proven in industry while utilizing the most cost effective, state-of-the-art technology available. The reliability of the overall control system was considered to be critical in view of the cost of downtime associated with the gold extraction plant. BRIEF PROCESS DESCRIPTION The dump leaching process treats approximately 1.2 million tons per year of low grade ore at an average grade of 0.035 oz/ton. Run of mine material is dumped on a lined leach pad and weak cyanide solution is applied by drip irrigation. Pregnant solution run off is pumped to carbon columns in the milling plant for gold recovery and the barren solution returned to the dump leach circuit. Average gold recovery is 60%. This process has little instrumentation and control associated with it. The milling operation treats 700,000 tons annually of higher grade ore (0.200 oz/ton initially, dropping to an average of 0.135 oz/ton after first two years). Recovery is directly related to head grade (fixed tail assay effect) and currently averages 96%. A single pass through a jaw crusher reduces run of mine ore to minus 12 inches. The ore is stockpiled and reclaimed by loader for grinding in a two-stage milling circuit consisting of a SAG mill and ball mill, the latter in closed circuit with hydrocyclones. Cyanide and lime are added into the SAG mill to start dissolution of gold as early as possible in the circuit. The ground product leaves the milling circuit at approximately 78% minus 200 mesh and is fed to an unique "double thickener" leaching-recovery circuit. This circuit has been discussed in detail in a paper by J. G. Mansanti et a1 (1). Two thickeners are arranged in counter- current configuration with three leach tanks. Overflow solution from the first thickener is treated by carbon-in-columns (CIC) for gold recovery with 85% of the soluble gold recovered onto this carbon. Underflow slurry from this thickener is pumped to the leach tanks, with a total retention time of 12 hours, and then gravitates to the No. 2 thickener. Overflow solution from the second thickener is used as a wash in the first thickener. Underflow slurry from the second thickener is treated in a carbon-in- pulp (CIP) scavenging circuit to recover the remaining 15% dissolved gold. Gold-loaded carbon from both the dump leach and milling circuits is stripped in batches using the Zadra hot caustic- cyanide elution process. Gold (and silver) is recovered from the hot strip solution by precipitation with zinc dust and the product recovered on Funda pressure filters. The precipitate is retorted to remove any mercury and then smelted into buttons. The buttons (approximately 80% gold, 15% silver) are shipped to an independent refiner in Salt Lake City, Utah, for further treatment.

Jan 1, 1990

By R. Jacobson

This paper presents comparison of proposed future solution and in-situ mining methods for in-situ leaching and recovery of uranium, gold, silver, and copper. We propose to show that in the future new extraction methods for each metal will eventually evolve for each hydrometallurgical process step, i.e. in-situ leaching, recovery, and concentration; precipitation and reduction; and drying, when required--without generating waste or bleed streams. In the future, to be highly profitable, these forms of mining will develop into an integrated system to eliminate waste streams using safe, powerful lixiviants. This integration will include, with some exceptions, improvements in the following areas of process development: 1. Lixiviant selection and in-situ leaching technique, 2. Target metal recovery and concentration (at a surface hydrometallurgical facility) from the circulating pregnant lixiviant stream, 3. Target metal precipitation or reduction, where decant water from the precipitate dewatering step is reusable for resin or solvent strip liquor (eluant), 4. Final preparation of target metal for sales, i.e. drying, purification, casting, packaging, 5. Restoration where most of the groundwater restoration process is performed at the same rate as multi-wellfield mining.

Jan 1, 1987

By F. N. Earll

4.1 – DEFINITIONS 4.1.1-MINERALS, ROCKS, ORES AND NONMETALLIC MINERALS F. N. EARLL A mineral is a naturally occurring inorganic element or compound. Most minerals are distinctly crystalline in habit, and many commonly occur in well-formed crystals or crystalline aggregates, but a few species are characteristically amorphous or noncrystalline. Nearly 2,000 mineral species are now recognized but, perhaps fortun¬ately, a rather large proportion are rarely encountered. Even so, the application of modern mineral-detection and identification techniques reveals the presence of surprising numbers of mineral species, even in deposits of outwardly simple mineralogy. A rock is an aggregate of grains or crystals of one or more minerals. Some rocks are essentially monomineralic. Thus, limestone consists largely of the mineral calcite, and many sandstones are dominantly quartz. Most rock varieties, however, contain a number of mineral species, and their classification or identification depends upon the minerals present, their relative abundance, and such factors as rock tex¬ture and origin. Ore is defined as an accumulation or deposit of one or more minerals that can be exploited at a profit. The profit criterion commonly is taken somewhat liberally, however, for, although always present, the specific mineral content that constitutes a viable product is transitory. The term ore, usually, but no longer consistently, is restricted to the description of mineral deposits that are of value for their metal content. Nonmetallic deposits, although similar to metallic ores in many respects, generally are valued for the minerals or compounds present, as with rock salt (halite), rather than for individual contained elements, though such minerals as elemental sulfur are of much value. 4.1.2-ROCK TYPES F. N. EARLL Rocks arc grouped tinder three broad headings: (1) igneous rocks formed by the crystallization of magma; (2) sedimentary rocks formed by surface processes of weathering and deposition; and (3) metamorphic rocks, which may have been derived from either of the foregoing, but have undergone important physical and/or chemical changes through the application of heat or pressure, or both. The three groups can he further subdivided for convenient reference. Thus, igneous rocks are described as being abyssal or plutonic if formed at great depth, hypabyssal if formed at intermediate depth, and extrusive or effusive if formed at or on the earth's surface. The sedimentary rock group is similarly subdivided into clastic sediments and chemical precipitates. The clastic sediments form by deposition of preexisting minerals that were released from their place of origin by weathering and were transported to their point of deposition as discrete grains

Jan 1, 1973

By T. J. Crebs

The latest estimates compiled by the Society of Exploration Geophysicists again indicate a moderate increase in mining geophysical activity in 1980 over the 1979 level. While North American activity remained at above the 1979 level, a considerable increase in mining geophysics was reported in South America, Australia, and the Far East. The total worldwide expenditures for mining geophysics were reported to be $53.7 million in 1980, compared to $44.4 million in 1979, and $31.6 million in 1978. During 1980, approximately 2% of all dollars spent on geophysics were attributed to mining geophysical activities; this percentage has remained relatively constant in recent years. Airborne surveys accounted for 51% of the total worldwide mining geophysical expenditure, 43% was spent for land surveys, and 6% for borehole surveys. Within the US, the breakdown of expenditures for land surveys was 60% for elec¬trical methods, 23% for gravity and magnetic methods, and 17% for seismic techniques. Electrical techniques remain the primary exploration tool for US mining geophysicists. Electrical Methods With inductive or electromagnetic (EM) techniques, significant developments were achieved in both frequency-domain and time-domain systems. Work continued on increasing the signal level on most time-domain (TEM) methods, to increase the exploration depth. The Crone group increased the power of its Pulse ElectroMagnetic system (PEM) to a 20-amp transmitter-loop capability. The GEOEX group is modifying the SIROTEM II system to obtain larger transmitter amperage from a portable motor generator. Geonics developed a new digital recording system (data logger) for its EM-37 system. This development should increase the productivity of EM-37 crews. A new ground, frequency-domain EM system was developed by the Scintrex group. This novel Genie system does not require a wire link between the receiver and transmitter. Because an amplitude-ratio is measured, the Genie data are reported to be relatively insensitive to coil orientation and distance errors. This new technique does not need extensive line-cutting or accurate station-chaining and would appear to be a good reconnaissance instrument. Scintrex also began marketing the new IPR-11 induced polarization spectral receiver. This receiver is microprocessor controlled, and can output to a cassette tape and record 10 windows of secondary voltage decay simultaneously from up to six receiver dipoles. The Phoenix group's new 100 kW induced polarization/resistivity (IP/R) transmitter began tests using their IPV-3 multifrequency, multichannel receiver. While this unit was primarily developed for "oilfield" IP exploration research, it has obvious application to "deep" mineral exploration. The Phoenix group also developed a new remote-reference, real-time magneto-telluric (MT) device in 1981. This five-component MT sys¬tem has a frequency range from 0.0005-384 Hz. Helicopter-borne electromagnetic (HEM) developments also continued in 1981. The mining in¬dustry increased its use of the new Geonics EM-33-3 multifrequency, multicoil instrument. In 1981, the Dighem group developed software for estimating magnetite as a mapping parameter from its HEM system. Dighem's work is said to complement airborne magnetic intensity surveys, since the HEM estimate is independent of remanent magnetism and magnetic latitude effects. Gravity and Magnetic Methods Probably one of the most innovative techniques in geophysics in 1981 was the use of airborne gravity surveys for both mining and petroleum exploration. The Carson group is using a modified, shipborne LaCoste-Romberg platform in helicopters. Data accuracies to 0.5 milligal have been achieved by flying gridded surveys. Although this airborne method is expensive-up to $186/ km ($300/line-mile)-the geophysical community has been excited by initial results. On the ground, the portable proton-precession magnetometers are becoming sophisticated. Both GeoMetrics and EDA recently introduced field magnetometers having data storage and processing capabilities. This development should greatly increase the productivity of ground-magnetic surveys. Seismic Methods Development of high-resolution seismic techniques continued in 1981. These techniques have primarily been directed toward coal studies for fault detection. OYO Instruments introduced their McSEIS-1500 seismic data acquisition system in 1981. This device contains a 24-channel recording capability, with digital output to 256-kbyte floppy disks. The high-speed data transfer using the disk media is considered a desirable feature. Borehole Methods The general decrease in uranium exploration, where borehole logging is extensively used, probbly led to the overall decline of geophysical logging activity in the minerals industry. However, a number of new sondes and logging systems were introduced in 1981: • Mount Sopris recently introduced their Series III logging system. This microprocessor-controlled unit records up to four channels of data on nine-track or cassette magnetic tape. The logging package is relatively light-weight, so helicopter transport to mountainous or roadless exploration sites is possible. (Both the Edcon and Woodware-Clyde consulting groups offer "slinging" capabilities for their Mount Sopris units.) Mount Sopris is continuing work on their 500-mm-diam (2¬in-diam) spectral gamma-ray sonde. This tool is expected to be available soon. • Owl Technical "slim-downed" its successful digital deviation probe to 380-mm (1.5-in) outside diameter. This new sonde will also measure inclinations up to 80° from the vertical, as compared with its older instrument that could measure inclinations to 30°. • A magnetic susceptibility sonde was introduced in the US by the OYO Instruments group. This Kappalog sonde contains two aircored coils for measurements slightly affected by thermal changes within the borehole. The increased activity in massive-sulfide exploration and the need to "look" deeper no

Jan 5, 1982