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By John W. Leonard

The typical tunnel boring machine in use on today's tunnel project uses 15'2-inch diameter cutters at a nominal load of 45 kips per cutter. The cutterhead rotates at a speed of 120 over D revolutions per minute where D is the cutterhead diameter in feet. The typical machine operates about 50 percent of the time. Delivery of a new machine takes 12 to 15 months. The TBM is a specialized tool and not applicable to all jobs. There is a range of job conditions that limit the economic use of a TBM. The machine should provide a thrust per cutter at least twice the compressive strength of rock in psi and up to three times the rock strength is required if in "granitic rocks." The rock strength in psi must be at least three times the cover above the tunnel in feet or the machine may stick. For example, 2,500 feet of cover requires at least 7,500 psi rock to avoid problems. When the RQD's over a significant distance of the project are less than 50, the TBM becomes support-bound. When a large percentage of the time is spent waiting while the ground is supported, the TBM cannot perform well (See Figure 1). This problem increased inversely with tunnel diameter. We are finding that increasing the RPM's does not significantly increase progress. The increased RPM results in a substantial decrease in cutter penetration. The explanation for this decrease appears to be vibration, cutterhead drag, or lack of stiffness in the cutterhead (See Figure 2).

Jan 1, 1981

By Fabiola Espinoza Carmona, Elena Chiriotti, Edoardo Misano

"INTRODUCTION Ganntas Tunnel is part of the modernization project of the railway between Alger and Oran, in Algeria. In order to double and rectify the existing line between El Affroun and Khémis Miliana, the alignment foresees the excavation of a 7 km-long twin tunnel. The excavation works started in June 2011 with the contractor CCECC, under the supervision of SYSTRA. Excavation is driven in conventional method by hydraulic hammer simultaneously on 8 different faces since excavation was started also from a junction window towards the middle of the tunnel. The minimum longitudinal distance to be respected between two contiguous tunnel faces has been set to 30 m. The tunnel cross-section is a 70 m² oval shaped profile, temporary support consists of shotcrete, bolts and steel ribs. A 30 to 50 cm thick cast in situ concrete final lining has been provided as well. When the tunnel reached a fault zone in soft clayey marls extreme squeezing occurred, works were stopped and re-profiling operations were carried out along more than 100 m. To date, works proceed at slow rate since high convergences are still monitored and completion of works is not expected before December 2016.GEOLOGY The project is located within the Tellian Atlas mountain area, including in particular the Djebel Ganntas massif in between Mitidja and Chleff plains. Djebel Ganntas is characterized by an extensive massif aligned in East-West direction with smooth slopes made of soft clays that emerge in multiple cliffs and ravines. On the top of it, along the edge, a more recent sandstone formation is found. The existing data on the geology did not allow to establish an accurate geological model at tender stage. In-situ investigations allowed at design stage to identify a Quaternary alluvial cover of clays and gravel, followed by a Miocene stratum divided in a sandy top layer and a lower clayey layer. From a structural point of view, geological maps suggest a north direction dip of layers, but the in situ surveys showed that layers were highly folded. The lack of information at tender stage on the structure of the geological series should have triggered an extensive investigation campaign at detailed design stage which did not take place causing serious consequences on costs and time at construction stage."

Jan 1, 2016

By L. Yuan, C. Brown, N. Rayyan, J. Soles, T. H. Dubaniewicz

As lithium-ion batteries (LIBs) become more prevalent in the mining industry, new hazards of battery fire and explosion are emerging. Efforts must be taken to ensure that workers are safe from these new hazards, such as batteries undergoing thermal runaway in underground areas that may have explosive methane-air mixtures. Researchers at the National Institute for Occupational Safety and Health (NIOSH) investigated overpressures generated within a sealed battery enclosure filled with an explosive methane-air mixture and a single cell lithium-ion battery driven into thermal runaway using an accelerating rate calorimeter (ARC). For both iron phosphate (LFP) and nickel manganese cobalt (NMC) lithium-ion cells, the explosion overpressures remained unchanged with varying percentages of methane concentration in the atmosphere surrounding the cell. It’s likely that the gases released from the battery undergoing thermal runaway caused an inert atmosphere within the sealed canister. The results from this study will help mining equipment manufacturers develop proper measures to keep miners safe while working with lithium-ion batteries in underground gassy mines

Mar 2, 2022

By P. G. Irwin

Natural resins, those derived from coal and those mined in the Uintah basin, face challenges from synthetic resins that are derived from petroleum sources and rosin or wood sources. In the past, fossil resins from coal were produced by various companies, the last one being Hercules, Inc., which sold BLACKHAWK resin until a fire destroyed the plant in the mid 1970's. Approximately four to fjve million pounds of BLACKHAWK were produced and sold mostly to the printing ink industry. Uintaite, more commonly known as Gilsonite® resin, was widely used by various industries, notably in printing inks, coatings, stains, adhesives, foundry, asphalt paving, and oil well drilling muds. Gilsonite has lost market share over the last thirty years to various synthetic resins particularly hydrocarbon resins that are derived from petroleum sources.

Jan 1, 1990

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION This method is a variation of cut-and-fill mining where either the wall rock or the vein is too unconsolidated to permit safe mining by conventional cut-and-fill methods. Basic development consists of driving a 2.4 x 2.7-m (8 x 9-ft) haulage heading laterally in the footwall side of the vein. Crosscuts are driven over to the vein at 91-m (300-ft) intervals. Then the vein, at each inter¬section, is mined out to a distance of about 9 m (30 ft ) above the track level. After the installation of an ore chute, scrams or intermediate drifts are driven in both directions along the vein. Mining then starts imme¬diately below the level bounding the upper limit of the stope. A cut varying between 3 and 4.2 m (10 and 14 ft) in depth and mined for a distance of 45.7 m (150 ft )along the strike is taken from the orepass in each direction (Fig. 1). In this mined-out void, the floor of the cut is equipped with 203 x 203-mm (8 x 8-in.) caps hitched into the wall with steel saddle hitches installed on 1.5-m (5-ft) centers. Double wood stringers, 152.4 x 152.4 mm (6 x 6 in.), are then installed on 1.2-m (4-ft) centers longitudinally down the length of the stope. Reinforc¬ing wire mesh is also placed, overlying each mat of timber. Directly above the stringers, stulls are installed from hanging wall to footwall to hold the timber in place during the sand-filling operation. Then a cement¬sand mixture in a ratio of about 1:10 weight is poured to a depth of about 2.4 m (8 ft). The entire cycle is then repeated down to the scram drift. Another variation of timbering is illustrated in Fig. 2. There is a vein width of approximately 5.5 m (18 ft) with the same unconsolidated ground conditions. The primary difference between handling this situation and the previously described undercut-and-fill method is that standard timbered sets are installed in each excavated stoping void rather than stulls. Once again the basic method is to work down the ore block from an upper timbered drift to a lower timbered drift. Cuts are taken from a timbered raise with a chute on the stope side. Each piece of ground excavated is essentially two sets high, 5.5 m (18 ft), by two sets wide, 5.5 m (18 ft), by the length of the cut along the strike of the vein. The timbered sets are 2.4 x 2.4 x 2.4 m (8 x 8 x 8 ft). The net result is, therefore, two sets of timber on a top floor separated by doubled-up 76-mm (3-in.) lagging from

Jan 1, 1982

By A. S. Winters

INTRODUCTION Homestake Mining Co. owns and operates the Bull¬dog Mountain mine near Creede, CO. Lead-silver con¬centrates are produced from ores extracted from narrow veins which dip from 1.04 to 1.39 rad (60 to 80°). An¬nual mill throughput is approximately 90 700 t (100,000 st). Ore shoots average 2 m (7 ft) in width and the ore is often weak, muddy, and poorly consolidated. Walls can be either strong or highly fractured along the strike of the veins. Cut-and-fill methods were initially selected to mine the deposit and considerable overhand mining occurred during the initial stages of the mine. Because of un¬stable ground conditions, mine management recognized that other methods would have to be employed. Undercut-and-fill mining was adapted to extract the loose ore from narrow veins and is presently the primary method used. Approximately 80% of the annual pro¬duction is extracted by this system. Loading ramps are positioned along the strike of the vein on 90-m (300-ft) intervals to allow for 45-m (150-ft) long cuts in each direction from the extraction raise. This interval was selected for efficient slusher operation and stope cycling. Loading ramps rather than chutes were selected for ore movement into 4.5-t (5-st) Granby-type cars because of extremely sticky ore conditions. Mined-out areas are filled with deslimed mill sands. Sand is pumped from the mill into two mine storage dams where the water is decanted. Each dam has a 544-t (600-st) capacity. Sand is pumped into the mine at a 30 to 40% slurry through 2195 m (7200 ft) of 51-mm (2-in.) thick wall pipe. When called for, sand is reslurried within the dams to a density of 55 to 65% and pumped to the particular mine opening requiring fill. Dry cement is added to the slurry prior to pumping. The capacity of the mine fill system is 20 t/h (22 stph). DEVELOPMENT Vertical development below the main adit level oc¬curs on 61-m (200-ft) intervals from an internal winze. Main haulage drifts or laterals are 2.7 x 2.7 m (9 x 9 ft) in section and driven parallel to the vein structures. Ground conditions generally dictate that better progress can be realized in the footwall than in the hanging wall. A distance of approximately 21 m (70 ft) is maintained between vein and drift openings to accommodate a smooth 18-m (60-ft) radius curve for crosscuts which are driven on 91-m (300-ft) intervals. Approximately 18 m (60 ft) of tail room is driven past the vein inter¬section to accommodate train loading. Ore and waste are hauled using diesel or electric driven locomotives operating over 18-kg (40-Ib) rail [762-mm (30-in.) gage]. Loading ramps are established above each crosscut¬ vein intersection. Scram drifts are then driven in the vein to determine the width and grade of the ore. If a prior decision to mine the block had been reached, then the scram is driven only far enough [about 12 m (40 ft)] to establish a drawpoint and scraper tail room. In most cases, the drawpoint is a 2.1 x 2.1-m (7 x 7-ft) cutout driven about 2.4 m (8 ft) into the footwall. On the level above, a raise bore station is established in the footwall to one side of the crosscut-vein inter¬section. The station is of sufficient size to accommo¬date a Robbins 41R raise boring machine. Support for the station is accomplished with 1.2 and 1.8-m (4 and 6-ft) roof bolts and wire screening. Once completed, a 152-mm (6-in.) concrete slab is poured and the raise drill set in position to center the 1.5-m (5-ft) diam hole approximately 1.8 m (6 ft) from the vein footwall. A 229-mm (9-in.) pilot hole is then drilled to the scram cutout below and reamed back to full diameter (see Fig. 1). STOPING Initial mining in a typical undercut-and-fill stope be¬gins from the top of the borehole. The initial cut can be either immediately beneath the track level or the old scram drift depending upon the particular situation. Cuts vary but are generally 2.4 to 3 m (8 to 10 ft) high. Support is accomplished by square-set timber, overhead caps, or roof bolts and screening depending upon ground conditions. Once the initial cut is removed, underhand timber and screen is installed and cemented sand fill placed. The stope is then ready for undercut mining. Prepara¬tion work begins with sinking 4.5 to 6 m (15 to 20 ft) on the borehole plus cutting a slot to the vein. As sinking progresses, hanging sets are installed to provide man¬way and service facilities. Hanging sets 1.7 m (5 ft 8 in.) long and 1.6 x 2.2 m (5 ft 4 in. x 7 ft 4 in.) in section are installed with steel hangers similar to shaft sets. Framed 203 x 203-mm (8 x 8-in.) timbers are made up in advance and shipped to the stope during sinking operations. When manway and service facilities are completed to the floor of cut to be mined, generally 3.65 to 4.3 m (12 to 14 ft) below the previous sand fill, the ore is removed to the hanging wall plus a round each way along the vein. Levelers or long stringers are then installed over the raise opening to support the work deck, slusher, and grizzly opening. All waste and ore from the sinking process is fed by gravity to the raise (see Fig. 2). Once the stope is equipped, mining pro¬gresses outward in each direction from the grizzly open¬ing. Stope equipment consists of 907 kg (2000 lb) pull air tuggers, timber skips, 11-kW (15-hp) ventilation fans, 15 to 22-kW (20 to 30-hp) electric slushers, 762 to 1270-mm (30 to 50-in.) scrapers, 67-mm (25/a-in. ) jackleg drills, as well as other drilling and blasting supplies. In narrow stopes [generally less than 2 m (7 ft) wide], no support is required for the underhand caps of the previous fill. Where wide sections in the vein occur, segment braces are installed beneath the caps to insure that the sand support cannot slip away from the hanging wall. Sections as great as 6 m (20 ft) wide and 4.5 m

Jan 1, 1982

By John A. Shuster, Joseph A. Sopko

Construction of two shafts, as well as associated deep concrete structures, for a major storm sewer project in Milwaukee, Wisconsin required ground water control and temporary ground support to depths ranging from 19.2m to 55.0m. Lowering of the site ground water table eras not possible because of potential consequences for existing nearby structures. Construction entailed completion of two shaft collars into bedrock. In addition to the shafts, several adjacent connected. structures were completed. Ground freezing was selected to provide temporary ground support and ground water control for all excavations. The first application of using reinforced frozen earth was successfully incorporated on the project. The procedure enabled the contractor to create openings in excess of 9.1m in height, without bracing, through highly stressed frozen earth walls. Design features and construction procedures for completion of ground freezing and related excavation are reported.

Jan 1, 1989

By Joseph F. Peters

The paper will begin with the physical parameters of design in a wire saw. These will include such things as the type of wire used, abrasives, horsepower of operation, and drive components. It will then go into the many uses of wire sawing in the quarrying and finishing of granite. This will include a step by step program of the removal of a block unit of granite from the quarry site through the finished slabbing and jointing in the plant. After this, there will be a small insertion of possibilities of wire sawing techniques in underground excavation work. This would be speculative only. I will use the example of powerhouse type of excavation work given to me by Dr. John Reed, and try to explain how I would attack this type of work using wire saws. Several slides of both design criteria and actual installations were shown.

Jan 1, 1997

By M. Stefanakis

Respect for the people is one of the main values of S&B. Such respect needs to be manifested by actions ensuring the bodily and spiritual integrity of personnel in pursuing their work. The ultimate target is to achieve zero accidents by developing the culture and infrastructure required to support it. Within this framework, the following proactive safety practices were introduced in 2004 and are implemented at S&B Greek sites: (a.) Management Safety Visits (MSVs) of shop floor operations to address safety attributes of work guidelines, (b.) Monitoring compliance with critical work guidelines, using a ?safety barometer? tool. The proactive nature of both above systems, facilitates communication with employees on safety behaviors and their adoption in our facilities at Milos and Fokis, has been successful. Above all, these practices corroborate the management?s concern for the people and its commitment to continuously improve the working environment.

Jan 1, 2007

By W. D. Keller

Scan electron micrographs, SEM's, will be shown of representative refractory clays from Missouri deposits, with counterparts from other localities. The story is in the pictures, ("1000 words in each picture"), but clay types include the following. Plastic, semi-plastic, and semi-flint "fire clays" of Missouri have textures which appear to retain certain relic texture of ball clay. Flint clay is composed of tightly interlocking packets, sheaves, and books of kaolinite. Diaspore and boehmite mineral textures characterize those "clays". Surface-boulder diaspore shows very pronounced recrystallization and crystal enlargement from original, unexposed first-grade diaspore.

Jan 1, 1977

By R. N. Torbin

Foster-Miller has recently developed a new concept.for a highly specialized shaft lining system. This innovative shaft lining system uses a continuous slipforming placement technique. lining the shaft from the bottom up such as in a raised bored shaft. Although this part is not unique, the concept is designed to use polyurethane foam instead of concrete for the liner material. Furthermore, the concept is designed to attain rates of advance of 30 m/hr. The polyurethane foam liner is designed to be thinner (100 mm) than a conventional concrete liner (300 mm). It will eliminate the high cost associated with the erection and operation of an on-site concrete batch plant. The system will be designed to be operated remotely, requiring few personnel for its operation. This system (or parts of it) may be appropriate where fast construction speed is required or where a jointless liner is needed.

Jan 1, 1988

By Rod Knopp

Each year approximately 5,000 tons of a sulfide concentrate grading 1 ounce gold per ton are produced at this operation. The material is refractory and as shipping costs and custom smelter charges increase, options to process it on site are becoming more attractive. Three methods of treating the ore have been investigated in detail. The first option of ultra fine grinding the concentrate produces a 35% gold recovery with relatively low capital and operating costs. A second option of pressure oxidation is very capital intensive and has high operating costs but recoveries in the order of 95% are achievable. The final option, biological leaching offers both capital and operating costs which are less than pressure oxidation, while still achieving 90% gold recoveries. Bioleaching of gold concentrates has been gaining acceptance but there is a risk factor associated with this or any emerging technology.

Jan 1, 1990

By Harry M. Castrantas

Caro's acid (peroxymonosulfuric acid-H2SO5) is used to destroy cyanides in tailings slurries due to its rapid reaction and low capital and operating costs. Due to its instability, Caro's acid is generated at the mine site using a hot or cold process. The majority of gold mines that use Caro's acid use the FMC hot process. In the FMC hot process, it is not necessary to cool the Caro's acid to minimize decomposition since it is used immediately upon generation. In contrast, the cold process uses a chiller to cool the raw materials and product. This paper will show, using laboratory and field data, that there is virtually no difference in Caro's acid yield between the cold and the FMC hot processes. Caro's acid yield depends on the mole ratio and concentration of sulfuric acid and hydrogen peroxide and is independent of operating temperature.

Jan 1, 1997

By Alfredo P. Vargas

Industrial flotation of sulfide ores is 80 years old. It has gone through many advances to date. However most flotation operations are still struggling to improve performance, which as a worldwide average is at the 90% level. What can be done to push the ceiling to 95% or better without performing finer primary grind, nor without duplicating flotation plant capacity? Such "unique task" can be performed by the low volume high capacity new flotation equipment known as the JAMESON CELL, when used as, "post-scavenger" of final plant tails. Improved efficiencies are obtained mainly from two cases: -Particles which never contacted a bubble. -Particles sheared-off by cell turbulence due to reduced number of hydrophobic sites. The application reverses conventional trend of using large volume mechanical cells by trade-in with smaller efficient equipment: resulting in reduced capital expenditures and operating costs. Outstanding "pre-rougher" applications are also presented.

Jan 1, 1993

By Brij M. Moudgil

Separation of oil shale from waste material or into high grade and low grade fractions can have a major influence on the economics of the surface (ex-situ) retorting process. Different components of shale rock vary significantly in physical and surface chemical properties such as specific gravity, color, water wettability etc. Various techniques which utilize the difference in one or more of the above propartis have been proposed to preconcentrate the oil shale before retorting. A discussion of the techniques proposed such as gravity separation, sorting and froth flotation is presented.

Jan 1, 1982

By Robert E. Ennis

Several important economic variables must be studied when planning a major underground conveyor haulage system if the objective of tow cost haulage is to be met. The purpose of this paper is to call these economic factors to the attention of the mine planner and encourage their serious consideration. To illustrate, the importance that some of these variables hole for the mine operator who wishes to reduce costs and maintain sound belt haulage practice, these areas of discussion will be covered. 1. Selection of economical flite lengths. 2. Economics of wire rope frame vs. rigid frame conveyors. 3. How idler design affects conveyor costs. 4. Economies in drive design. 5. How speed affects conveyor cost. 6. P.V.C. vs. Neoprene belt - what savings? 7. Maintaining, conveyor investment through maintenance. 8. Miscellaneous factors affecting the cost of belt haulage.

Jan 1, 1967

By Mahmoud Sepehrmanesh, Verya Nasri

The purpose of this paper is to outline the advantages and disadvantages of single tube and twin tubes for subway systems. Different aspects such as diameter of the tunnel (tunnel spaceproofing), configuration of subway line and stations, impacts on passenger safety and evacuation, ventilation system as well as rail system for each solution were evaluated and compared. In addition, a multi-criterion approach is established to facilitate the decision making for these two different subway systems. The practical example of a mega project has been discussed and the details of the decision-making process has been explained. This proposed multicriterion can be used to facilitate selecting the best subway system in any urban areas.

Dec 1, 2022

By J. W. Merks

Spreadsheet software is an effective tool for simulating dynamic stochastic systems such as mineral processing plants. The two-product formula is used to show how to set up a Monte Carlo simulation in a spreadsheet template and how to calculate variances of dependent variables. The key to a reliable and realistic simulation model is to obtain unbiased estimates for all independent variables and variances in the set that defines the stochastic system. Monte Carlo simulation is also a powerful tool for optimizing sampling procedures, for applying sensitivity analysis and for designing and evaluating experiments.

Jan 1, 2000

By W. F. Riggs

INTRODUCTION Flotation equipment is becoming more definitive in design as well as greater in dimension. The last ten years have witnessed the implementation of large flotation cells ranging in size from 28 to 85 cubic meters. However, small plants still operate with cells as small as 3 cubic meters. These changes have increased the operational sensitivity of the flotation system to the chemical reagents used. No scientific study has been undertaken to develop the relationship between the size of equipment and the response of reagents in the flotation system. Hence, this article is an attempt to present several observations from past experience for the purpose of stimulating future research in an area previously ignored and perhaps subsequently not used to advantage. CONVENTIONAL FLOTATION Recent advances in conventional flotation cell design have been made in two areas: a. Flotation cells are becoming larger, and b. Basic design of flotation cells has changed. They can be classified into two types: deep cell and shallow draft cell. Size and design changes of flotation cells have a greater effect on frother performance than on the performance of other chemical reagents. Hence, the discussion in the following sections is restricted to the impact of frothers in the flotation system. Flotation Cell Size The advent of increasing flotation cell size has resulted in the increase of frother consumption. There appears to be a correlation between reduced horsepower (kilowatt) consumption and increased frother consumption per ton of ore processed in the circuit. This is more prevalent when the frother chemistry remains unchanged. Another reason for increased frother consumption is the increased distance between the rotating central shaft and the lip of the cell for larger flotation machines. Increasing this distance results in increased surface area corresponding to the larger cell size. Reduced power input per ton combined with the need for longer, stronger load support of the froth bubble as a result of increasing cell size has resulted in the development of a new series of stronger and more persistent frothers. By the same token, these newer, stronger frothers are difficult to control when applied to the smaller cells. This is true even at lower dosages because the break point between too much frother and not enough frother for optimum metallurgy is normally very narrow. This problem is often exaggerated when operators have established an optimum setting for pulp level and air for the particular frother in use. Generally, the tendency is to operate the flotation cells simply by adjusting frother dosage. The increase in frother strength to accommodate the large cells often results in the realization, after a retrofit, that the existing pump box or sump which is still handling the same volume of concentrate is no longer adequate in size. Flotation Cell Design The two most common flotation cell designs installed during the past ten years are the deep cell and the shallow draft cell. The deep cell provides for complete mixing of air and pulp at the bottom of the cell. The shallow draft cell does all the mixing/pumping in the bottom haft of the cell whereas the mixing of the air is performed further up the shaft and is brought into contact with the pulp midway up the vortex. There are many preferences for one or the other. It is the author's contention that both designs have very valid concepts and must simply be suited to the particular situation as required by the operator. The following comments are subjective observations which should be explored in greater detail. The deep cell mixes both air and pulp together at the bottom of the cell requiring a full depth ascending of the homogenized air to the surface to form the froth layer. This design, in combination with charged air, provides for good concentrate grade control under adverse conditions of selectivity, especially when faced with a fine grind and a slime-laden gangue. A selective frother can be applied without a competitive loss in recovery. One of the conditions for good selectivity in this environment is operating at medium pulp solids concentration in the range of twenty-six to thirty-four per cent (26-34%). When a higher solids concentration is used in combination with a coarse grind, it becomes difficult to maintain the same recovery with a selective frother due to higher pulp viscosity. This condition sometimes results in an unacceptable coalescence of the air bubbles as they ascend the full height of the deep cell. The increased pulp viscosity at higher solids concentration can also retard the flotation kinetics. The combination of decreased surface area of dispersed air and slower kinetics can result in reduced recovery causing coarse particles to continue to drop

Jan 1, 1989

By Deepak Malhotra

The first step in the development of a gold deposit is the detailed mineralogical examination of the ore types, major and minor mineral constituents and their association. It is extremely difficult and expensive mineralogically to characterize the occurrence and association of gold with other minerals for low-grade refractory ores (less than 0.1 troy oz/t Au). Even if a mineralogist can identify a gold-associated particle or two, one wonders how representative they are of the deposit. What alterative does a metallurgist have for characterization of gold-bearing ores? Diagnostic leach procedures, consisting of a series of sequential leaches, can provide the metallurgist/mineralogist with reasonably good Indication of the deportation of gold in the various mineral matrices. This information con help the metallurgist concentrate on the most promising process options In the Initial stages of the project development process.

Jan 1, 1993