Search Documents

By L. Xu, N. I. Jayaraman, R. Bhaskar

The objective of this study is to develop a better understanding of the collection characteristics of dust measurement instruments used in mine dust control research with respect to properties of quartz and coal dust. Coal and quartz dust of narrow sues, with aerodynamic median diameters being approximately 2 µm, 4 µm and 10 µm, with quartz contents of l0%, 20%, 30%, 40% and 50%, were released in an aerosol characterization chamber using a fluidized bed aerosol generator. Dust was collected in MSA personal gravimetric samplers with l0-mm nylon cyclones, total dust samplers and Sierra 298 inertial cascade impactors. The samples were analyzed for mass and quartz content. A total of 21 experiments were performed for this study. The mass and quartz collection efficiencies of the total and respirable dust samplers and the impactor were determined. The data show that the amount of dust collected by the respirable dust sampler (respirable + oversize dust trapped in the cyclone pre-classifier) was, in general, less than the amount collected by the total dust sampler. The amount of dust collected by the impactor was higher than that collected by the total dust sampler. Quartz analyses of the airborne dust samples showed that the total dust and respirable dust samplers (including oversize dust) collect less quartz dust than at inertial impactor. The larger the median particle sire of the quartz and coal mixture, the lower the quart content of the dust collected in the samplers. For all three sires, the impactor was a better collector of quartz than the other instruments.

Jan 1, 1994

By A. C. Lyons

Foreword Sir William Halcrow and Partners are responsible for the design of the Ahmed Hamdi Tunnel as Sub-contractor to the Contractor, Osmac, a joint venture formed by the Arab Contractors, Osman Ahmed Osman and Company of Cairo, and Tarmac International of London. Throughout the paper the term "Client" refers to the Employer, the Ministry of Housing and Reconstruction, which latterly became the Ministry of Development and New Communities of the Arab Republic of Egypt. At the time of writing of this paper only a few hundred metres of the tunnel have been driven, but the design is substantially complete. Introduction The Ahmed Hamdi Tunnel, which is at present under construction, will be a two-lane tunnel crossing beneath the Suez Canal approximately 15 km north of the town of Suez at El Shallufa. It will become the first permanent connection across the Canal since 1967 and is visualised by the Egyptian Government as an early step in the development of the Sinai Desert along the eastern bank. The tunnel will carry pipelines for treated water and electric power cables to assist in this development. The tunnel, which will be 1640 m in length, is being driven through hard clay and mudstone using a tunnel shield fitted with hydraulic excavators. The primary lining is being constructed with precast, bolted reinforced concrete segments with an internal diameter of 10.4 m. The road deck structure and the walls of the air supply ducts beneath it are formed with precast, reinforced concrete units erected immediately behind the shield.

Jan 1, 1979

By Peter J. Tarkoy, Terence G. McCusker

INTRODUCTION The discovery of oil and gas reserves in the North American Arctic has led to substantial thought being given to their development. Systems adequate for exploration are not always, or even often, suitable for permanent installations. Combinations of offshore pipeline, drilling ships, platforms, sub-sea systems, sand islands and -ice island bases have been proposed. However, conditions in the Arctic are such that these methods may be difficult or impossible to implement. Problems exist with (1) the freezing of the Arctic Ocean surface for 9 - 10 months per year, (2) the movement of sea ice, (3) scouring of bottom sediments by ice islands and icebergs to great depths below the ocean bottom and (4) deep water channels. Any method which can avoid the problems associated with surface construction in these very adverse conditions merits detailed examination. One of the proposed solutions is to tunnel from a convenient or available access point to a point within the field, construct a chamber large enough to accommodate the drilling equipment and drill the development wells in a protected environment. Service and delivery lines would use the access tunnel and further tunnels and chambers would be constructed in a radiating pattern to develop the whole field. In order to determine the feasibility of underground construction in the North American Arctic, it is first necessary to understand the geologic and environmental factors which would affect such construction, to be knowledgeable concerning the problems and then to consider the techniques appropriate to the conditions which will be encountered, bearing in mind the need to minimize both con-

Jan 1, 1976

By S. J. Bluhm

INTRODUCTION Lesotho is a mountainous area of southern Africa from which water is to be exported in an extensive tunnel system, to industrial regions inland. The related tunnelling project has involved a num- ber of drives using tunnel boring machines [TBMs] to excavate about 100 km of 5 m diameter water tunnels [von Glehn and Bluhm, 1995). This paper describes the ventilation and cooling of some of the tunnel drives from both the operational and design points-of-view with a particular focus on heat generation. There were many common features in all of the drives but this paper is focused mainly on the Hlotse drive which was 18,4 km long. The drives were ventilated using forced ventilation systems to provide appropriate air flow throughout the tunnels and face zones. In addition, the Hlotse drive required refrigeration equip- ment which provided chilled water to the tunnel. While the sec- ondary ventilation systems play an important role in gas and dust handling, the paper concentrates on the primary ventilation and cooling issues. The ventilation of these tunnels was an exacting exercise be- cause: • Rock temperatures and geothermal heat flow were high. • TBMs with relatively high power ratings were used. • Diesel locomotives were used. • Drives were relatively long. • High altitude meant a low air density. An important feature was the simulation and monitoring of the ventilation and heat flow components and the project was characterised by analysis, monitoring and ongoing tactical decision-making throughout the progress. The thermodynamics of the systems were complex because there were many interactive effects and analyses were carried out using special computer pro- grams. The monitoring confirmed the accuracy of the models, and in this manner it was possible to confidently ensure healthy and safe working conditions and still minimise costs. Local ambient climate conditions range from temperatures higher than 35 "C in summer to below -10 OC in winter. Based on available statistical data and the thermal storage/damping effects in the system, design summer ambient conditions were taken as 15 OC/25 "C wet-bulb/dry-bulb. The barometric pressure was 80 kPa and due to the altitude, the ambient air density was only 0,9 kg/m3. The local Authority specified a maximum in-tunnel wet- bulb temperature [at any point] of 32,O OC and a mean wet-bulb temperature [from all locations] of 27,5 OC maximum. The maxi- mum height of ground cover above the tunnel was 1 200 m and the maximum virgin rock temperature was 41 OC; see Figure 1. Diesel dilution criteria specified by the local Authority was a minimum of 0.1 m3/s per rated kW of diesel engine. Other requirements related to gases such as CO, CO2, NOx and CH4 [and the need for intrinsically safe equipment] but these are not of direct relevance to this paper. The actual average face advance was about 30 m/d with good days achieving 60 m/d and good months achieving 1 000 m [23 working days]. The original design tunnelling rate was 50 m/d. DESCRIPTION OF HLOTSE DRIVE VENTILATION AND COOLING SYSTEM The ventilation requirements in the tunnels were dictated by heat and diesel dilution needs. The best ventilation and cooling policy is generally a balance between using increased quantities of fresh air or refrigeration [or both]. In this particular scenario it turned out that, since the diesel emission criteria required large quantities of air, the refrigeration needs were modest. The drive was ventilated using a ducted, forced ventilation system from fans located at the portal. The maximum ventilation requirement was 51 m3/s when the drive was at 18.4 km. From a heat flow point of view, the worst scenario was a heat load of 3.5 MW when the drive was at 7 km. This was cooled by the ventilation air and a supply of chilled water to the tunnel. Refrigeration and chilled water system In the design phase, a detailed comparison was carried out between two general alternatives for providing refrigeration. First, was a system in which refrigeration sets and air coolers are installed on the TBM train; the refrigeration sets are cooled by condenser water piped to and from cooling towers at the portal. Second, was a system in which refrigeration water chillers are in- stalled at the portal and chilled water is piped into the tunnel. The detailed comparison indicated that the capital and running costs of the second system were at least 60 % lower than the in-tunnel plant. There were also many obvious practical benefits for favouring the portal system. The refrigeration plant supplied 23 11s of cold water at a temperature of 10 OC. After providing the cooling effect in the drive, the water returned to the portal where it was initially cooled in open-circuit evaporative pre-cooling towers, chilled in the refrigeration plant and then returned to the tunnel. The cold water flowed into the tunnel in an insulated supply pipe and returned in an uninsulated pipe; the water was simply circulated to the end of the pipe and returned. The cooling effect in the tunnel was achieved entirely through heat transfer from the pipe [long linear heat exchanger] and no air coils or other heat exchangers were required. The cooling requirements were satisfied by the heat transfer to the returns chilled water steel pipe [200 mm]. The pipes were eventually installed to a maximum distance of 10,8 km in what was considered a very practical and cost effective solution.

Jan 1, 1997

By R. F. Holub, E. Franco, T. B. Borak, K. J. Schiager

INTRODUCTION According to a recent analysis by Schiager et al.(1981), inhalation of radon daughters continues to contribute the largest risk from exposure to radioactive materials in underground uranium mines. The American National Standards Institute (1973) has published a concensus standard for radiation protection in uranium mines. According to this standard, a monitoring system for radon daughters must be capable of measuring the annual accumulated exposure in working level months within an uncertainty interval of 50% at the 95% confidence level. To satisfy this criterion, uranium mining companies in the USA have adopted a procedure consisting of single or periodic measurements of the air concentration of radon daughters. All of the measurement methods sample a known volume of air through a filter. After a specified delay interval, the filter is analyzed for radioactivity. The exposure rate from total airborne potential alpha energy in units of working level (WL) is estimated from the measured activity on the filter. For control purposes, this measurement establishes whether or not a work area is safe for occupancy by miners. For routine dosimetry, the concentration estimated from the grab sample is assigned to the appropriate work area. It is then assumed that the concentration of radon progeny is equal to the average of the two measurements at the beginning and end of the specified interval. Meanwhile, the work locations and exposure times of each miner are recorded on the daily time card and verified by the shift supervisor. Most cards provide space for 15-minute time intervals. The exposure in working-level hours (WL-hr) for each shift is obtained by summing the time spent in a given location multiplied by the exposure rate (WL) assigned to that location.The exposure of an individual miner to radon progeny is summarized on a weekly basis and commonly reviewed each month. These values are then accumulated in units of working level- months (WLM) during the calendar year. EVALUATION TECHNIQUE AND RESULTS The evaluation of grab sampling methods for monitoring exposure to radon progeny is presented in two sections. The first section examines each method individually according to accuracy, precision, and uncertainties attributed to human errors. The second section includes an analysis of procedures which are common to all measurement methods. The ultimate goal is to determine whether a specific measurement method, or grab sampling in general, can fulfill the criterion that the annual accumulated exposure of each miner be known to within ± 50% at the 95% confidence level. In addition it is essential that the above criterion be satisfied at an acceptable cost to the mining company. The results of each testing procedure are presented in terms of the contribution of the relative uncertainty in the estimates of cumulative WLM. Each of these are then combined to form a figure of merit for the complete monitoring system. The method of combining statistical uncertainties and interpretation of the results is subject to some debate. However, since all of the tested parameters enter into the final computation of WLM multiplicatively, we have used the commonly accepted procedure of combining the squares of the relative uncertainties. We have also assumed that each of the uncertainties is derived from independent random variables and the covariance of any two variables is zero. A description of each grab sampling measuring method is given in Table 1. Also included is the equation used to compute WL from the measured quantities. Accuracy (U A) Our definition of accuracy is based on the uncertainty inherent in any method which attempts to estimate potential alpha energy without unambiguously determining the concentration of the three unknown radioisotopes, RaA, RaB and RaC. (For simplicity, the RaC and RaC' nuclides are referred to in combination as RaC.) The inherent uncertainty strongly depends on the mixture of these three components at the time of measurement.

Jan 1, 1981

By Charles R. Nelson

INTRODUCTION The support of openings in weak sandstones can be achieved by spraying on a liquid grout which will soak in and harden to form a shell. In the St. Peter sandstone of the Minneapolis-St. Paul area of Minnesota compressive strength is increased from about 0.7 MPa (100 psi) to over 7 MPa (1000 psi) to depths of 152 mm (6 in.). A shell of this strength and thickness can pro¬vide support much as shotcrete does but at lower cost. If the spray-grouted shell is inadequate for the support needed, it provides a good base for shotcrete application, which is often difficult in weak or friable sand¬stones because the fresh shotcrete peels off along with a thin layer of sandstone. The spray-grouted technique was initially developed at the Civil and Mineral Engineering Dept. of the University of Minnesota using funds provided by RANN Div. of the National Science Foundation. It was used as temporary support in a 3-m (10-ft) wide, 3700-m (12,000-ft) long storm water drain tunnel by the Minneapolis Sewer Construction Dept. Much of the field development at this construction project was carried out with the cooperation of the personnel of the city of Minneapolis. It also was used for temporary support in a 3.3-m (11-ft) wide, 1800-m (6000-ft) long storm water tunnel built in 1978-1980 for the Minnesota Dept. of Transportation in Minneapolis. In 1975, 2000 m (6500 ft) of 1.5-m (5-ft) wide utility tunnels were spray-grouted for final lining. A 50-m (150-ft) long, 1.2-m (4-ft) wide test tunnel built at the university in 1974 was sprayed for half its length and is being observed for long-term behavior which to date (1981) is good. GROUT MATERIALS The requirements for the liquid grout are more severe than those for injection grouting. Besides being able to penetrate the rock and develop sufficient strength, it must have the following properties: (1) be nontoxic, noncombustible, and have a low odor for spraying under¬ground; (2) have controllable viscosity and setting time for adjusting to the permeability of the sandstone and the required depth of penetration; and (3) must "wet" the sandstone and develop capillary "draw" to penetrate to the required depth. A sodium silicate-based grout with the proper setting agent meets these requirements in the local St. Peter sandstone which has a uniform grain size dis¬tribution curve with D,,, = 0.1 mm, permeability of about 10 to 20 darcys, a porosity of about 25%, and less than 5% of silt size or smaller. It is 97 to 99% pure SiO2. The grout mix consists of Philadelphia Quartz Co. Type N sodium silicate and Celtite 55 Terraset Spray Grade (SG) distributed by Celtite Inc., Cleveland, OH. A volume mix ratio of 100:9:125 for sodium silicate to setting agent to water produces a grout with two to four centipoise viscosity, and initial set at 20°C of 20 min. Setting time is temperature dependent, being longer at lower temperatures, but can be adjusted by varying the setting agent concentration a few percentage points. APPLICATION The application is simple. Spray the grout on the surface until the desired penetration is achieved. Penetration rates of about 5 mm (0.2 in.) per min. are realized in the St. Peter sandstone for up to 152 mm (6 in.) of penetration (30 min. of spraying). The set time must be longer than the spraying time. The spray nozzle should be moved back and forth at a rate that minimizes runoff due to surface buildup. Both a single large hand-propelled nozzle and multiple small machine propelled nozzles have been used successfully. The pumping and mixing equipment used consisted of: (1) a hand-held, hand-pumped, 11-L (3-gal) gar¬den sprayer for test patches (mix by shaking); (2) 76-L (20 gal) batch mixing in barrels with small electric pumps for spraying and mixing; (3) two component pumps consisting of Hypro Model 5300 piston pumps (Hypro Pump Co., St. Paul, MN) on a common shaft driven by an electric motor. The sodium silicate is premixed with a portion of the water and the setting agent is mixed with the rest of the water. The two components are combined and mixed just before the nozzle; and (4) a proportioner metering the setting agent just before the nozzle with the water and sodium silicate premixed and pumped from the surface. The proportioner permits the bulk of the grout to be pumped through a single pipe or hose with only the setting agent stored underground. This is efficient and low in cost. COST The grout mix costs about $0.20/L ($0.75 per gal) in 1978 dollars. Filling all the voids 100 mm (4 in.) deep (25% initial porosity) would have a materials cost of about $5.40/m2 ($0.50 per sq ft). The labor cost of application is usually less than the material cost for normal tunnel jobs. A pump and proportioner cost less than $1000. Sodium silicate storage tanks and pip¬ing costs would depend on the particular site conditions. REFERENCES AND BIBLIOGRAPHY Nelson, C. R., 1977, "Spray Grouting for Tunnel Support and Lining," Underground Space, Vol. 1, No. 3, pp. 241¬ 245. Yardley, D. H., Nelson, C. R., Stocker, T. H., 1974, "So¬dium Silicate Spray Impregnation of Tunnels in the St. Peter Sandstone," Research Report, University of Min¬nesota, 118 pp.

Jan 1, 1982

By Ralph S. Mattson, Frederick T. Hancock

GENERAL DESCRIPTION The Granduc mine is situated on the west side of Granduc Mountain in the Leduc River area of north¬west British Columbia about 51 km (32 miles) north¬east of the town of Stewart. Entry is by means of a 16.6-km (10.3-mile) long haulage adit from the concentrator site at Tide Lake. The principal ore mineral is chalcopyrite. Ore mineralization varies from dis¬seminated to irregular semimassive stringers generally associated with highly folded and sheared biotite-rich sections of metasediments. The ore bodies vary in width from 3.6 m (12 ft) to over 18 m (60 ft) in a zone some 800 m (2600 ft) in length dipping at an average of 75° to the west and plunging steeply to the south. Capping varies in thickness from 0 to over 300 m (1000 ft). SELECTION OF MINING METHOD In the early stages of planning the primary require¬ments to be met by the mining method were: (1) suit¬ability to high output and productivity, (2) adaptability to mechanization in order to minimize operating costs, and (3) minimization of the costs and the development period for preproduction. Considering these, the selection of a mining method was narrowed to a choice be¬tween sublevel long-hole stoping or sublevel caving. A method utilizing backfill was ruled out as uneconomical, especially since mill tailings would not be readily available. At the time, the following facts about the ore body were recognized: (1) the average width of the ore body was 12 m (40 ft), (2) 30% of the stoping blocks were in areas of 6 m (20 ft) or less in width; (3) in many cases, ore zones were located in parallel lenses or were irregular in shape; (4) major faults were present in the hanging wall close to the ore, with minor faults within the different ore zones; and (5) the ore itself was lami¬nated with fractures and jointing crossing such bedding planes. Considering these factors, the two stoping meth¬ods were compared. Sublevel Open Stoping The characteristics of sublevel open stoping are: 1) It is generally not used for widths below 6 m (20 ft) since the ratio of waste development increases as level intervals are reduced to insure good control of long-hole drilling. 2) This method is used where ground is relatively competent. Otherwise more pillar support is required, increasing ore losses or mining costs with pillar recovery. 3) The method is not easily adaptable to parallel ore zones separated by relatively narrow waste bands. 4) Close control of pillar and stope design will per¬mit some degree of dilution control. Sublevel Caving Sublevel caving is characterized by the following: 1) The method can be used for a variety of ore widths, recognizing that 3 m (10 ft) in width may be an economical minimum. 2) It is particularly favorable where wall rock is weak. 3) The method has flexibility so that irregular ore bodies or parallel lenses can be mined. 4) The need to leave pillars for support is eliminated 5) Dilution and tonnage recovery are interrelated. With concern about the stability of the hanging wall and the geometry of the ore bodies, there was a growing favor to choose the sublevel caving method. However, this method was relatively new to Canadian mining practice. Teams were sent to Sweden to tour mines using sublevel caving and to study their methods and any extraction and dilution problems. The objective was to develop cost and performance data and ore recovery grades and tonnages for feasibility comparisons. Next, detailed mine plans were developed covering a three-year period using sublevel caving and sublevel open stoping. Extensive analyses of the development and production costs for each method were made. With sublevel caving, ore losses of 10% with a 20% dilution rate were used. The same ore loss was estimated for sublevel open stoping with dilution varying from 11 to 15%. As a result of these studies, the decision was made to use sublevel caving partly because slightly lower op¬erating costs, together with higher productivity, offset the adverse effects of higher dilution; and partly be¬cause, at the start, there would be less risk in being able to handle the ground, and the method offered more flexi¬bility. Open stoping was not ruled out for testing at a later date. It was recognized that some other method of mining would be required for mining the ore extensions below the glacier, where surface subsidence must be avoided. During the operating life of the mine, open stoping was tested in narrow width ore zones and found un¬satisfactory due to wall failure. A cut-and-fill stope, utilizing run-of-mine waste from development, was also operated on a trial basis. A sequence was finally estab¬lished that permitted ground control; however, produc¬tivity needed considerable improvement in order for open stoping to be considered as a sole alternative for mining below the main haulage level. At the termination of operations in June 1978 over 13 000 000 t (14,500,000 st) had been extracted from above the main haulage level at production rates varying from 3270 to 7250 t/d (4000 to 8000 stpd). Sublevel Caving Methods Two basic methods were used, transverse and longi¬tudinal. The transverse method was used mainly in the upper C ore body in widths greater than 18 m (60 ft), and occasionally as an alternative to the longitudinal method in localized areas where strike faulting inter¬fered with normal development. The longitudinal method was used in ore bodies varying in width from 3.6 to 18 m (12 to 60 ft), with a multilongitudinal varia

Jan 1, 1982

By G. W. Riley, D. E. Pickett

HINDERED SETTLING CONCENTRATION In the free settling of mineral particles in a liquid, the falling particles are at a distance from each other so that no particle is affected by its neighbor. In hindered settling, the concentration of particles is sufficiently high so that each particle is affected by its proximity to other particles in the suspension. Richards and Locke86 have described the hindered settling phenomenon as the condition “.. where particles of mixed sizes, shapes and densities in a crowded mass, yet free to move along themselves, are sorted in a rising current of water, the velocity of which is much less than the free-falling velocity of the particles but yet fast enough so the particles are in motion." This is the condition normally encountered in mineral con¬centration processes. The well known Newton equation for free settling of coarse (ap proximately +10-mesh, 1/16-in. or -- 2000-µm) spherical particles is:87 where v," is free settling velocity, cm/sec; p' is density of the fluid; p is density of the particle, g/cm3; d is particle diameter, cm; g is acceleration due to gravity, cm/sec2; and Q is coefficient of resistance, dimensionless, ~0.4. For the settling of fine spheres in water (approximately 150 mesh or 100 µm) the equation of Stokes pertains:87 where µ is viscosity of the fluid in poises and the other symbols have the same meaning as those in Eq. 1. For particles whose size lies between about 10 mesh (--2000 µm) and 150 mesh (100 µm), their settling velocity can be determined from experimental data. These data are available in convenient form in the text by Taggart88 based on the original work of Richards. Alternatively, a Reynolds number-coefficient of resistance plot may be used to determine the settling rate of such particles.87 The settling rate of spherical particles under hindered settling conditions can also be calculated from Eqs. I and 2 by replacing p', the density of the fluid, by p" the apparent density of the suspen¬sion. The concentration of particles in the fluid thus imparts an appar¬ent density to the composite fluid or suspension greater than that of the liquid alone, resulting in a buoyant effect on the larger particles. Particle shape affects the settling rate of both coarse and fine particles. The general effect is to reduce their settling velocities and the effect is greater for coarse particles and for those settling under hindered settling conditions than for fine particles or free settling ones. For two particles of differing densities but settling at the same velocity under Newtonian conditions, the ratio of their diameters from Eq. 1, called the free settling ratio is: where L signifies the lighter particle and H, the heavier particle. Under Stokesian conditions the exponent would be 0.5. For hindered settling conditions the fluid density p' is replaced by the apparent density of the suspension, p", to obtain a generalized equation for the hindered settling ratio: assuming both particles settle in approximately the same regime. The free settling ratio as given by Eq. 3 has been called by Taggart88 the "concentration criterion" and is used to predict the effectiveness of any gravity concentration process (see Introduction to this section). Based on Eqs. 3 and 4, if two particles of densities pH and p,, settle at the same velocity, the diameter of the lighter particle will be larger than that of the heavier particle. For example, in the case of galena (pH = 7.5) and quartz (pL = 2.65) settling in water (p =1.0) under free settling, Newtonian conditions 3.9. Thus, a quartz particle nearly four times as large as a galena particle will settle at the same velocity. Any quartz particle just slightly less than four times the diameter of the largest galena particle may be separated from it. Under hindered settling, Newtonian condi¬tions in a suspension where p" = 1.65, dL/dH = 5.85 or any quartz particle just slightly less than about six times the largest galena particle may be separated from it. Reference to Eq. 4 indicates that a superior separation between two minerals of differing densities is favored by: (1) coarse particles settling under Newtonian conditions, (2) a large difference in (pa - PL), and (3) separation under hindered settling conditions where p" is high. Of course, there are practical limits to increasing p" excessively because at very high percent solids suspen¬sion fluidity would be lost and the hindered settling separation process defeated. Examples of hindered settling separators are the Dorrco-Fahren¬wald sizer,89 the Rheolaveur box 90 the Spitzkasten,89 and the Willoughby washer.91 These devices make a mineral separation on the basis of both specific gravity and size and all of them are essentially obsolete except for the Dorrco-Fahrenwald sizer and similar devices which still find application for the removal of coarse particles from a much finer particle assemblage and for preconcentration ahead of shaking tables. However, nearly all gravity concentration processes (jigs, tables, flowing film concentrators, heavy media separators) and many sizing devices (sizing classifiers, clarifiers, thickeners, hydrosepa¬rators) make use of the hindered settling phenomenon during the separation of particles. JIGGING Introduction In jigging, a mixture of ore particles, supported on a perforated plate or screen in a layer or "bed" with a depth many times the thickness of the largest particle, is subjected to an alternating rising and falling (pulsating) flow of fluid with the objective of causing all

Jan 1, 1985

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores, such as cobalt, platinum, gold, silver, etc. Flotation of metal sulfide ores has been successfully practiced at the industrial scale for more than 50 years. It is the relative technical/economic ease of floating metal sulfide minerals that has been a major driving force behind the rapid expansion in the supply/demand of metals over this 50 year period. It is estimated that 1.6 billion tons of feed sulfide ores are processed each year in the free world based on the froth flotation process, Klimpel (1987) and the manufacture/use of associated chemical reagents is a major portion of the 1.5 billion dollars (U.S.) free world mining chemicals business, Klimpel (1988b). The use of froth flotation appears to be increasing partially due to the ever decreasing feed grades and liberation sizes of the feed minerals. Froth flotation has shown itself to be a flexible process that lends itself well to many solid/solid separations. Thus, with surprisingly little equipment modification, separations can involve very different particle sizes and densities, and relative weight ratios of materials to be separated. In addition, the flotation process is economic, especially when compared to size reduction, its associated precursor process. Flotation also lends itself to continuous operations with a variety of equipment configurations. Such operations exhibit an ability to vary feed rate of solid to the process by as much as 50% without a total collapse of separating efficiency. In addition, the widely used mechanical flotation cell is scalable from 2.8 to 57 or 85 m3 (100 to 2000 or 3000 cu ft) with surprising ease relative to other unit engineering operations over the same relative size increase. The separating medium used is water. Most of the chemicals required -- pH regulators, frothers, collectors, activators, and depressants -- are all relatively inexpensive and common. They are not usually used in large quantities. While there has been a tendency for the quality of metal sulfide ores in any one geographic location to deteriorate over time due to mining intensity, the overall global supply of sulfide minerals is still very ample. Thus, the physical supply side of sulfide ore concentrates has not been under strong technology improvement pressure with increasing metal demands. Part of this is due to the ability of capital to consistently move to new geographic sources having suitable-grade sulfide ores. In these moves, the actual froth flotation technology used (including collector chemicals) has had to change little. As an illustration of this trend, of the approximately 80,000 metric tons of thio collectors used commercially (1980) in the free world, almost 987 were known and manufactured in some form 25 years ago. This is clearly not typical of reagent development in other process-related industries. In addition, the industrial-scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully on an industrial scale for many years. Thus the industrial process of froth flotation, especially as applied to the recovery of sulfide minerals, poses a dilemma: such a process is widely used, economic, versatile, forgiving, etc. and yet is still not very well understood in a mechanistic sense leading to prediction of results from fundamentals even after 50 years of usage. Lack of such predictive ability has not limited the general over-all industrial use of froth flotation in sulfide mineral recovery. Rather the lack of

Jan 1, 1989

By David R. Klug, Chris Gause, Troy L. Page

The Inter-Island Tunnel is one of two undersea tunnels which are part of the Boston Harbor Clean-Up Program. The tunnel is 7.7 km (25,160 ft) long, bored to a 4.3 m (14 ft) diameter in the Cambridge Argillite Formation and was finished to a 3.5 m (11.5 ft) diameter final concrete lining. The owner is the Massachusetts Water Resources Authority (MWRA) who is a regional authority responsible for overall design and construction. The tunnel's primary purpose is the transport of untreated effluent from Nut Island facility in Quincy, MA to the new treatment plant at Deer Island. Two sludge lines are installed for the transport of sludge from Deer Island to Nut Island. During construction the contractor encountered fault zones, fractured strata, and soft ground tunnel sections which liberated more ground water than was originally anticipated by the project designer. Ground water was controlled during mining and the placing of the final lining. The final lining of cast-in-place concrete had to be backfill grouted to cut off water inflows and provide contact for equal tunnel loading which is critical in long-term design. This paper deals with an engineered backfill grouting system which was the first of its type to be used in North America. It discusses the geological conditions encountered, the design criteria used in the evaluation process, and an indepth review of the various system components. It demonstrates how contractors can use knowledgeable and professional industry suppliers to solve difficult construction problems.

Jan 1, 1999

By Joseph R. Hatch

Previous studies of the Thompson and Temperly ore bodies show alteration aureoles in the Middle Ordovician host-rock limestones, dolomites and shales extending laterally as much as 60m from the ore. Thickness, texture, carbonate and clay mineralogy, major-, minor-, and trace-element contents; and carbonate carbon and carbonate oxygen Isotope compositions of the host rocks all have been significantly altered by hydrothermal processes. Organic geochemical analyses of samples from the basal 8-cm-thick carbonaceous shale bed and immediately overlying limestone bed of the Quimbys Mill Member of the Platteville Formation, collected from within the ore, the alteration aureole, and the unaltered country rock show: 1) solution of carbonates and residual concentration of organic matter, 2) decreases in the volatile (Si) and pyrolytic (S2) hydrocarbon yields In samples of the basal shale from within the ore, 3) decreases in the relative amounts of the lower molecular weight (<nC25) saturated hydrocarbons in samples from the ore, and 4) no changes in either the organic carbon isotope composition or the Rock-Eval or biomarker thermal maturation Indicators. In light of the 125-170°C depositional temperatures documented for the early marmatitic sphalerltes in these deposits by fluid Inclusion studies, the lack of thermal alteration of the organic matter suggests a relatively short mineralization episode, a conclusion supported by the relatively simple sphalerite band stratigraphy. The decreases in the volatile and pyrolytic hydrocarbon yields for shale samples from the ore supports a model for mineral deposition which obtains sulfide through reaction of brine sulfate and indigenous organic matter.

Jan 1, 2013

By Luiz Roberto Amaral, Peter Frobenius

Construction of the first phase of the Sao Paulo Metro, the North-South Line, will be completed late in 1974. The downtown section of this line consists of 2.2 miles of 19.6-foot diameter soft ground tunnels and 3 subway stations. The entire tunnel - station package was let to one contractor. A plan and profile of the tunnels and stations is shown on Figure 1. Passage of the tunnels under Boa Vista Street in the Sao Paulo Financial District was one of the major challenges. The street is only 50 feet wide from building line to building line and flanked by 10 to 20 story buildings. The tunnels pass within 4 feet of adjacent footings. Further south, the tunnels run 13 feet below the flat foundations of the Caixa Economica, a massive 11 to 14 story building. Surface settlement had to he limited to less than .07 feet/tunnel and good face control and grouting procedures were essential. For this reason, full face cutting wheel tunneling machines were selected to drive the tunnels south of the Access Shaft Prestes Maia which included Boa Vista Street and the Caixa Economica. The tunnels north of the Access Shaft Prestes Maia were driven under a fairly wide Avenue and greater settlements were permitted. Open face shields were selected to drive these tunnels. Most of the work has now been completed and analyzed. Here are some of the results - Careful planning and execution of tunneling procedures kept surface settlement over the tunnels below a magnitude of .05 feet, allowing tunneling adjacent

Jan 1, 1974

By M. S. Celik

Despite its importance in various diverse industries, research on beneficiation strategies of boron minerals has not received enough attention. Boron minerals are primarily upgraded by wet processing methods. However, severe environmental constrains have accelerated the development of dry processing techniques such as calcination. Interestingly, boron minerals, colemanite, ulexite and borax undergo respectively decrepitation (fragmentation), shrinkage and expansion upon heat treatment. Each mineral then requires a different beneficiation scheme. In this study, laboratory and pilot plant studies on calcination of boron minerals conducted in our laboratories are reviewed along with their implications in commercial operations.

Jan 1, 1997

By Ryne Goodnough

INTRODUCTION The Wyodak Mine is located in the northern Powder River Basin of Wyoming and is thought to be the oldest continuously operating coal mine in the United States, having recorded yearly production since 1923. The mine has expanded over its 80-plus-year history from an underground room and pillar operation to a modem surface snip mine. Currently, the mine feeds four on-site power plants using conveyors, has a filly automated truck load-out, and ships coal by rail to nearby Wyoming power plants. Annual production of 4.8 MTFT is accomplished on a single-shift basis with 53 fill-time production, maintenance and administrative personnel. In 1990 a decision was made to modernize the coal feed system to the power plant complex with modular conveying equipment. The project involved constructing coal storage silos, overland conveyors, feeder breakers and modular conveyors to follow the mining benches. Total project cost in excess of $25 million was used to construe or commission: A 2,150-TPH 48-inch overland conveyor with horizontal curve A 2,150-TPH 48-inch plant feed conveyor 13 shuttle conveyors, each 150 feet in length 2 front-end loaders equipped with large coal buckets 2 crawler-mounted feeder breakers with compatible discharge conveyor 1 high-lift shuttle conveyor with adjustable discharge height 1 high-angle conveyor 2 4,500-ton coal silos with automated truck load- out capabilities. Conceptually, the plan was to develop 33 MTONS of reserve adjacent to the plant complex and use front-end loaders to tram ROM coal to in-pit feeder breakers. The primary crushing would reduce coal to 6-inch minus and feed through a series of shuttle conveyors to an overland supply conveyor. The overland conveyor would feed the secondary crusher, and the crushed product would be moved to any of four storage silos using conveyors. The in-pit crushing and conveying system is a unique fit for the power plants and provided many interesting challenges for the coal mine to overcome. System integration, maintenance windows, equipment availability and the like had a major role in engineering the system to a power plant complex having no emergency coal stock- pile and limited plant storage. In crediting the designers and operators, the in-pit conveying system at the Wyodak Mine has delivered coal to the power generators since 1994 with little or no downtime. ENGINEERING CONSIDERATIONS Yearly production for the mine is based on 3.5 MTPY which requires about 1,550 TPH on an 8 hour, one-shift- per-day, 5-days-per-week basis. Considering efficiencies for starting and stopping, crew transport, loader trams, and operational delays, a nominal conveyor feed rate of 2,150 TPH or 72% system utilization was used in engineering design application. This ton-per-hour basis is used to size and commission the remaining components of the in-pit system. Coal quality requirements also played a role in equipment selection. The mine reserve grades vertically and coal from Upper and Lower benches needed to be mixed in any given proportion. This required equipment to be filly variable and capable of producing all or part of the 2,150-ton-per-hour requirement. Bisecting the 80-foot- high coal face for quality required construction of a high lift module capable of reaching a single overland conveyor 40 to 50 feet above the pit floor. Designers built a high-lift, adjustable and portable module specific for this application Having two bench systems capable of each producing the required feed for the power plants required computerized control processes to control on-line production. RSLinx, RSLogix, and RSView were chosen as the communication and control software, and a Data Highway Plus serves as the communication backbone. Each of the two in-pit benches are controlled with a data center equipped with Allen Bradley 5-40 processors. This

Jan 1, 2004

By Ricardo D. Presnell, Erich U. Petersen, William J. Tafuri, Donald M. Hausen, Russell C. Babcock, William L. Gunter, Thomas R. Gibson, Douglas N. Halbe, Jay W. Hammitt

The Barneys Canyon orebody is a sediment-hosted disseminated gold deposit located 8 km (5 miles) north of the Bingham Canyon porphyry copper mine. Mineable oxide reserves are 8.6 million tonnes (9.5 million short tons) grading 1.64 g/tonne (0.048 oz/ton) gold. Host lithologies are carbonates and siliciclastics of the Park City and Kirkman-Diamond Creek Formations of Lower Permian age. Gold distributions are controlled by favorable stratigraphic units and associated alteration, and structures. Dominant alteration is decalcification and bleaching of carbonates, and argillization and bleaching of sandstones. Gold to silver ratios are greater than 1:1. Anomalous trace metals include arsenic, antimony, mercury and thallium. The Melco deposit lies 2.4 km (1.5 miles) southwest of the Barneys Canyon deposit. It is spatially associated with a steeply northwest-dipping siliciclastic breccia in the Kirkman-Diamond Creek Formation at the contact with underlying orthoquartzites of the Freeman Peak Formation. Both gold-bearing oxidized and sulfide lithologies are present, often in sharp contact. Mineable oxide reserves are 2.8 million tonnes (3.1 million short tons) grading 2.40 g/tonne (0.07 oz/ton) gold. Anomolous trace metals include arsenic, antimony, mercury and thallium, which are generally present in concentrations greater than those at Barneys Canyon, particularly in sulfide-bearing material. A spatial, temporal and genetic relationship is proposed between the Melco deposit and late-stage gold mineralization at Bingham Canyon, although both the Melco and Barneys Canyon deposits have many similarities to Carlin-type gold deposits.

Jan 1, 1990

By J. Oliver, J. Cleason

"With maintenance costs rising and an increasing demand for higher production, the use of sealed spherical roller bearings can reduce maintenance costs while providing increased production uptime. The sealed spherical bearing has increased bearing life by slowing the ingress of contamination. This paper will describe how contamination and bearing wear is reduced by reviewing case histories from coal and metal mines that improved their production by decreasing downtime using sealed spherical roller bearings. INTRODUCTION The spherical roller bearing (SRB) is a robust rolling element bearing that can accommodate high radial loads, low to moderate axial loads, and misalignment with no loss in carrying capacity. (i.e. Figure 1) Because of these characteristics, it is a favorite in mining applications and many other heavy industry applications. Primary applications are the various pulleys on conveyors. Coal and hard rock mines have hundreds of miles of conveyor lines and countless head, tail, drive, and idler pulleys. Maintaining the bearings on these pulleys and replacing bearings that have failed can be a huge expense to a mine in components, labor and downtime. Ideally, the bearings are replaced when the pulley is changed out according to the maintenance cycle, about every 4 years. Reality, however, causes the bearings to be changed out every year, or sooner. These premature bearing replacements can cause production shut downs for as long as 8 hours. Conveyor pulleys have been targeted at many mines as the single biggest maintenance pain point.The primary reason for replacing conveyor bearings is internal wear from fine, solid particle and water ingress through the bearing housing seals. (i.e. Figures 2 & 3) Traditionally, the bearing is mounted in a split pillow block housing (SAF, FSAF, SDAF or SNL are typical industry styles) with only labyrinth seals (LER/LOR) for protection against the ingress of contaminants. The bearings are lubricated with grease. A common way to extend the MTBR is to purge the bearing system with large quantities of grease to flush out the contaminants. If the bearing, and thus the housing seals are not frequently purged with grease, fine solid contaminants enter the housing and bearing through the labyrinth cavities."

Jan 1, 2015

By R. E. Vivas, J. A. Villa

"INTRODUCTION Located in the state of Colima, Peña Colorada is one of the largest iron ore operations in Mexico. The project is owned in equal parts by ArcelorMittal and Ternium. Though boasting a production capacity of 4.5 million tons of concentrate per year, Peña Colorada is faced with declining ore grades. In order to respond to this situation, Peña Colorada had to expand production even further. This expansion created various challenges that were overcome by the adoption and application of new technology. Examples of these technological innovations are discussed in detail in this paper and include improvements in geo-modeling, mine planning and mine operations. The successful implementation of these technologies paid for itself many times over, and has led to significant improvements in production and performance, while lowering the operating costs and reducing the variance between planning and execution. PROJECT OVERVIEW Peña Colorada is a mining company which produces about 30% of the iron ore in Mexico. This project generates about 1,200 direct jobs and 3,000 indirect jobs, making a significant contribution to the local economy. Similarly, Peña Colorada invests in its local communities through numerous educational programs, scholarships, university agreements, social development, environmental agreements, etc. Peña Colorada’s operations consist of an open pit mine located in Minatitlán and a pellet plant located in the port city of Manzanillo, as shown in the location map below:TECHNOLOGICAL INNOVATIONS Most mining projects around the world typically follow the logic of “mining the next best ore”. Consequently, head grades tend to decline over the life of the mine. In the case of Peña Colorada, a decline in head grade combined with a decline in iron ore prices forced the operation to adapt to a new business environment and adopt new technologies to face these challenges. However, the first step along this process was to increase the confidence and reliability of the resource model. This task involved additional drillhole campaigns coupled with more detailed modeling techniques. Once the block model was completed, new optimization studies and production schedules were evaluated with new optimization tools that unlocked significant project value in many aspects of the mine plan. Similarly, new geotechnical studies were prepared, which provided new guidelines for pit slope design, and new technology was implemented to monitor ground movement and slope stability. The picture below shows two sectional block model views in MineSight 3D, along with the pit designs and geologic solids"

Jan 1, 2017

By Lynn I. Kelley

Industrial-mineral production contributes on the order of CDN$1.5 billion per year to the economy of Saskatchewan, Canada. Potash is the primary industrial mineral produced in the province, followed by aggregate, sodium sulphate, salt, potassium sulphate, peat, clays, silica sand, calcium-chloride brine, and "clinker." Potentially economic deposits of kaolin, diamond, and building stone are under evaluation. All of the major deposits of commodities discussed here are located south of lat 56°30&apos;N. Most of the locations are presented on Figure 1 POTASH Saskatchewan supplies approximately 30% of the world&apos;s demand for potash. Eight conventional underground mines and two solution mines produced an all-time high of more than 13 million metric tonnes of potash (KC1) (8 million tonnes K2O equivalent) in 1997, valued at over CDN$1.4 billion. All of the mines exploit the middle Devonian Prairie Evaporite. The four potash-bearing members of the Prairie Evaporite underlie much of southern Saskatchewan and extend southward into North Dakota and Montana and east- ward into Manitoba (Fig. 2). Halite and anhydrite separate the potassic members. The evaporite beds locally are interrupted by solution structures (anomalous ground). Some of the structures may be related to deep-seated tectonic features.

Jan 1, 1999

By M. G. Ayat, B. C. Scott

H. P. Ferri I refer to the article "Degradation Process in Coal Slurry Pipelines" by M.G. Ayat and B.C. Scott in the September 1988 issue of Mining Engineering. I would like to congratulate Dr. Ayat and his colleague on a very interesting and valuable piece of work. This article is an excellent one, describing a real phenomenon that has been ignored for so long. I do, however, have an objection. The article fails to address the degradation process when coal is processed or transported in other ways than slurry form, particularly those employed in this work. Could, for example, screening (wet or dry) have contributed to the extent of degradation? Perhaps scanning electron micrographs of the samples examined and after screening would have helped to clarify this. I would like to point out that degradation is also observed in pneumatic conveying systems, mechanical conveying, and other means of transportations employed by the mineral industry. Reply by M. G. Ayat We gratefully acknowledge Dr. Fern&apos;s comments on the subject paper. As he so rightly points out, the degradation process is a real problem, not only in coal slurry pipelines, but also in many handling and processing equipment employed by the coal and mineral industry. Scanning Electron Microscopic examination of samples may offer very little information about the extent of degradation for several reasons. One is that preparing a representative sample that has undergone degradation will be extremely difficult if not impossible. This problem is exacerbated by the fact that SEM requires only very small amounts of material for testing. Secondly, there is a sample preparation step involved in the SEM examination and, as such, it would be very difficult to associate the extent of degradation to either sample preparation or handling and processing operations. Finally, it is probably worth mentioning that degradation is a multi-faceted phenomenon involving many natural and man-made factors. The logical approach to this problem would, therefore, seem to be a "one-step-at-a-time" approach.

Jan 2, 1989

By Dale C. Matthews

Pressure oxidation of refractory gold ores is becoming more popular in our environmentally sensitive society. This paper will discuss some specific process problems and solutions involved with starting up and operating an autoclave circuit.

Jan 1, 1991