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By C. O. Aksoy

There is a long history of coal production in Turkey with electric energy generated from power plants since the 1950s. However, the cost of power generation is high. Therefore, it has become necessary to find a means of decreasing production costs to become more competitive in a global market. This paper attempts to economically minimize the overall operating costs of a dragline-shovel-truck system applied at an open pit mine that has fed the Yatagan power plant since 1979, lowering the lignite production cost, and indirectly, the overall electricity generation unit cost. Systems currently in place, as a result of former optimization studies, are time-consuming and expensive to re-implement to comply with today’s very different geometric and economic conditions affecting mining operations. In such cases, rather than re-optimization, it is more viable to identify the parameters that affect the minimum unit cost, via a cost minimization study, allowing present mine equipment to be utilized more efficiently. Unit cost minimization analyses are subject to the capacity of the pool of mining equipment available, combined with the mining block geometry from pre-stripping onwards. In this overall unit cost minimization study, P&H shovels and CAT 777 trucks with capacities of 15 yd3 and 77 tons, respectively, were used in truck-shovel pre-stripping operations. A 65 yd3 Marion 8050 dragline was used to expose the overburden immediately over the coal seam. Pre-stripping bench heights varied from 8 m to 20 m and dragline bench heights from 20 m and 25 m were evaluated to minimize the unit cost of field operations. Total overburden extended to a depth of 60 m (35 m to 40 m for pre-stripping). Minimization analyses performed on the pre-stripping truck-shovel operation, prior to dragline operations, determined that the shovel bench height should be 15 m requiring a dragline bench height of 25 m.

Jan 1, 2005

Taking the Presidential chair of the Australasian Institute of Mining Engineers, my first duty is to thank you for the high honour and position you have conferred on me. A position entailing the highest compliment the Institute could bestow on any of its members, and one which I feel my inability to properly fulfil, inasmuch, as you have men in your ranks of the highest scientific training residing on that great continent where the head quarters of the Institute is established, while my engagements here will not permit me leaving this colony but for a very limited period.About six years ago, during the time the mining boom was at its height, there came to this colony from England, Germany, and South Africa, several mining engineers, who in conjunction with the mining engineers here, formed an Institute which we soon found we were not able to maintain. The mining boom gradually died away, and many engineers have left the colony to find employment in other parts of the world. We made overtures to your Iqstitute to take us under your wing, which you were kind enough to do, and the outcome is that you have come over 1200 miles to hold a meeting in New Zealand. The paucity of members residing in this colony does not admit of a large attendance; but we heartily welcome you to our shores and hope you will enjoy a holiday visiting some of the principal mining centres and exchanging ideas, which not only improves our minds, and widens our field research, but also tends to lead to further economic improvements in the extraction of metals the earth contains...

Jan 1, 1903

"The Ernest Henry Mining (EHM) ore body is an iron oxide copper gold deposit with an average grade of 1.1 percent copper, 0.5 g/tonne gold and 22 percent magnetite. At the end of 2009, EHM converted the ball mill media charge from forged steel to high chrome media. At high pH values (above 10.5), forged media is passivated, so as the pH of the system drops the media corrosion rate increases and floods the system with corrosion products, affecting flotation. As the high chrome media is much more inert, pH has less of an effect on copper flotation response. Given this, the pH in the roughers was dropped from 10.6 to 9.6 and it was found that there was a slight increase in copper recovery and no statistical change in either the rougher or final copper concentrate grades for copper and gold. With the reduction in pH, there was a noticeable reduction in lime usage from 0.85 kg of lime per tonne of ore processed to 0.55 kg/t. This reduction in lime is equivalent to approximately $700,000 per annum at current production rates.INTRODUCTION The Xstrata Copper Ernest Henry Mine (EHM) is situated 38 km north east of Cloncurry, Australia. The mine began as an open-cut operation in 1998 and transitioned to a multi-sourced operation in December of 2011, utilising ore from the newly completed underground mine in addition to ore mined from the Mount Margaret Mining operations. EHM produces a combined copper gold concentrate containing 25 000 t of copper and 35 000 oz of gold annually, utilising a three stage grinding circuit along with a four stage flotation circuit. In addition to this EHM also produces 1 000 000 t of magnetite concentrate annually via a further grinding and three stage magnetic separation circuit. Based on metallurgical test work performed by Magotteaux between 2002 and 2008, the step to change from forged steel media to 18 percent chrome media was taken in 2009 in a bid to alter the pulp chemistry to improve the flotation performance in the plant (Greet, Kinal, Twomey, & Peatey, 2010). This plant trial resulted in: • Increased dissolved oxygen levels within the pulp, • A ball mill discharge Eh shift of 25 mV to more oxidising pulp potentials, and • An EDTA extractable iron percentage reduction. These improvements in the pulp chemistry resulted in an increase of 1.7 ± 0.6 percent in copper recovery and an increase of 2.9 ± 1.7 percent in gold recovery. As the plant trial using high chrome media was successful, the use of the media was adopted as the standard for the ball mill. This implementation opened the door for further process improvements around the flotation circuit."

Jan 1, 2014

By Stuart Brashear, Robert Brush, Ben Cook

In early 1993, the Piney River quarry owned and operated by the Blue Ridge Stone Corporation of W W? Boxley received a series of complaints from the owners of a 130 year-old historic farmhouse that had been converted into a bed and breakfast establishment. Since the distance from the quarry to the farmhouse was substantial, a unique method of seismic data acquisition was utilized to determine if perceived blast effects were caused by air blast or low amplitude, low frequency ground vibration.

Jan 1, 1996

Baja Mining is currently constructing the Boleo metallurgical plant in Santa Rosalia, Mexico to recover copper, cobalt, zinc and possibly manganese products from the Boleo ore. The process involves an acid seawater leach of the whole ore followed by solid-liquid separation and washing of the leach residue. Solvent extraction is used to separate copper, cobalt and zinc from solution. Copper and cobalt are recovered as cathode metal products and zinc as a zinc sulphate monohydrate. Manganese may be directly precipitated as manganese carbonate which can be used as a feedstock for manganese metal production. Laboratory testing has continued to refine various aspects of the process including iron removal, cobalt and zinc solvent extraction, cadmium cementation, cobalt electrowinning and manganese metal electrowinning. The results of this testing are reported in this paper.

Jan 1, 2012

By Mohamed Yusuff

Modern transportation systems are one of the most important achievements of our civilization and provide us with tools for broadening individual and social freedom of choice. The transportation system initiated today is handed dawn to the next generation, probably with little choice because of the enormous expenditure of monies involved, and influence their way of life as to where they should live and where they should work. The ever increasing antagonism between the citizens and the planners of the transportation systems necessitates a careful review of the entire planning process. The planning process can be broadly classified into: the goal formulation and decision making process; the plan development process; and the plan implementation process. THE GOAL FORMULATION AND DECISION MAKING PROCESS Goal formulation and decision making is one of the vital phases of the planning process and is primarily based upon social values. In measuring social values, the planner must employ an appropriate scale of measurement since in the absence of this yardstick, these values tend to became vague and abstract ideals. The need for major social programs and concerns (poverty, etc.) of the urban areas should be analyzed to determine the type of transportation system which would contribute to the solution.

Jan 1, 1974

By J. R. Donald

For processing copper, continuous converting has been successfully employed in a variety of locations, but for nickel, there has yet to be a successful commercial implementation of continuous converting technology. At Inco?s Copper Cliff smelter, there are specific requirements for the production of Bessemer matte relating to downstream processing that makes continuous converting for nickel mattes even more challenging. The nickel carbonyl refineries require very low iron levels in the matte, and the sulphur deficiency must be carefully controlled such that the appropriate quantity of PGM-rich metallics are produced from matte separation. The present paper outlines continuous converter technologies that have been implemented in copper, and evaluates their potential for nickel converting. The challenges of producing a Bessemer matte to the required specifications are discussed. The economic and environmental driving forces, and the challenges of justifying a novel technology in place of productive existing equipment, are discussed.

Jan 1, 2005

Impala Platinum?s No.16 Shaft mine is currently under construction and is expected to reach full production of 256 500 tons per month by 2014. Mining will start at 1 260 m below surface with a maximum mining depth of 1 535 m where virgin rock temperatures of 56.6°C will be experienced. As a result of depth, high virgin rock temperature and narrow reef mining, an expected total of 31.5 MW[R] of air cooling capacity will be required to achieve acceptable underground conditions. This will be achieved by cooling mine intake air in a single dedicated air cooling installation located on surface. The compact installation includes air cooling spray chamber, refrigeration machines, condenser cooling towers, evaporator and condenser water reticulation as well as all electrical and control systems. Construction and commissioning of the first phase of this air cooling installation has recently been completed on a turn-key contract basis.

Jan 1, 2009

By Karla Kaboth, Brad Garraway

"The surface of many concentrate thickeners has traditionally looked like that of a flotation cell. Over the last three years Normandy Golden Grove Concentrator, in partnership with Outokumpu Technology have developed and implemented an innovative breakthrough deaeration technology to address this problem.When installed in the thickener feedline, the system successfully detached air bubbles adhering to the solid particles and removed air bubbles from the feed stream, subsequently presenting a fully deaerated feed to the thickener feedwell.Tests to date have shown an 80% reduction in overflow solids. As a result, costs associated with concentrate loss and measures to recover concentrate that settles in process tanks and ponds have been minimized.MOTIVATIONIt is recognised in mineral processing operations that the entrainment of finely divided air in thickener feeds can cause significant problems in the operation of a thickener. This is particularly the case with thickening of flotation concentrates, in which air bubbles are strongly attached to the mineral particles and are difficult to remove. Such particles attached to fine air bubbles rise to the thickener surface, and hence over the weir and into the thickener overflow. High solids carryover can lead to a direct loss of mineral product, and if the product settles in process water tanks and ponds, the recovery of this product by means of intermittent dredging operations is usually expensive and inconvenient.Many concentrate thickeners are fitted with devices which attempt to alleviate this problem. Froth rings or baffles, skimmers and froth sprays, all play a part in reducing the amount of solids carried over into the overflow. However, many froths are very stable and persist despite these measures. In the worst cases, a thick froth layer forms which builds up to a depth greater than the froth ring, and hence may overflow or pass beneath the ring . Any concentrate that finds its way past the froth ring can also cause blockage of the overflow weir and launder, and has the potential to upset clarification performance by cutting off sections of the overflow handling system. The formation of a froth layer on the thickener also makes it more difficult for the operator to monitor the clarification performance."

Jan 1, 2001

By Gustav Wirth

"ABSTRACTIn exothermic and endothermic catalytic reactions, control of the temperature is of utmost importance. Basic calculations are developed for determining catalyst volumes of various types of reactors that give particular temperature distributions, and the individual designs are compared.INTRODUCTIONA narrow range of permissible operating temperatures characterizes many, catalytic reactions, and yields of the desired product are reduced sharply if operating temperatures fall outside of this optimum range. In order to achieve close temperature control, reactors with internal heat exchange have been developed in which thin layers of catalyst are located between cooling surfaces.The temperature distribution for an exothermic reaction in such a thin catalyst layer is shown in figure 1. Reactive gases flow downward through the catalyst section, which is 2so wide. It is assumed that a constant amount of heat is generated per unit of bulk catalyst volume, and that this heat is entirely removed through the walls. Consequently, except for a short section at the entrance of the reactor, the heat of reaction is not available for raising the temperature of the gases. Thus, at all points equidistant from the axis, the temperature is the same throughout the entire reactor. At the middle of the gas stream, the temperature is designated as uo. Traveling radially from this point, the temperature decreases because of the removal of heat through the walls and reaches, in the vicinity of the cooling surface, the value u1. Over a relatively narrow zone of width, b, the temperature falls to the value u2 at the cooling surface. Because inadequate data are available on heat-transfer coefficients for a gas flowing through a porous bed, and also because the narrow zone, b, is probably without effect on the reaction by virtue of its low temperature, the temperature drop in this zone at the cooling wall was disregarded. The following calculations were made on the basis of the temperature drop (u0 - u1) resulting from heat conduction within the main body of the catalyst bed only."

Nov 1, 1950

By S. J. Oosthuizen

This article explores the Fray Farthing Chen (FFC) Cambridge process, a novel method for the electrochemical deoxidation of metal oxides in molten salt, discovered at the University of Cambridge in 1997. The process was hailed as a highly promising, potentially low cost, novel method for the production of titanium metal direct from its oxides. The article should inform researchers in the field of some of the challenges in the commercialization of a novel, high profile process involving multiple stakeholders. The author, former senior process engineer at British Titanium Plc, the company originally tasked with commercializing titanium production via the FFC Cambridge process, reviews the latest literature and discusses past and present progress in the pursuit of low cost titanium metal via this process. Topics explored include the history of the process, attempts at commercialization, NASA?s alternative application, and present status of the process. Keywords FFC Cambridge, titanium, molten salt.

Jan 1, 2011

By P. Roy

Smelting or converting of sulphide concentrates and mattes in a batch operation presents a challenge when directing the off- gas to a conventional acid plant. The Hindustan Zinc Ltd. (a division of the Vedanta Group of companies) Chanderia Lead Zinc smelter was presented with such a challenge that was met with the adoption of the Cansolv® SO2 Scrubbing System. The Ausmelt top submerged lance technology used at Hindustan Zinc generates a furnace gas stream that contains between 0.1 vol% and 12 vol% SO2 during the three stages of batch smelting. Cansolv® SO2 Scrubbing Technology captures the SO2 to ensure that treated flue gas exits the facility at a concentration below 100 ppmv SO2 and it releases SO2 in a controlled fashion, to more closely match the needs of the existing acid plant. The application of a regenerable scrubbing system to capture SO2 from a batch-operated furnace and convert it to acid in a conventional acid plant was a first of its kind in the world. This paper provides an update on the design and operation of the Hindustan Zinc SO2 recovery system and suggests two improvements to the basic flow sheet that optimize the cost of SO2 scrubbing further.

Jan 1, 2007

By J. P. D. Strydom

The transmission of excessive vibration to the human body can cause physical stress that may result in crippling pain or permanent disability. Rock drills are known to transmit high levels of vibration that are difficult to attenuate through conventional approaches to vibration isolation. However, it can be shown that once the ISO frequency weighting has been applied, most of the vibratory effect of the rock drill?s motion is concentrated at the drill?s operating frequency. This observation opens the way for utilizing a concept known as vibration absorption, aimed at attenuating the rock drill?s vibration. In essence this entails using the inertia of a secondary vibratory system to counteract the motion of the handle. This article documents an investigation into the use of such an absorber for attenuating rock drill vibration and demonstrates the feasibility of such an approach. A mathematical model of such an absorbing handle was developed and used for design studies. Based on these investigations an experimental prototype model was manufactured and tested. A significant reduction in vibration levels could be achieved.

Jan 1, 2002

By J. L. Green

Virginia-Carolina Chemical Corp., in its Florida Operations, uses both roller mills and ball mills in dry grinding of phosphate rock. The roller mills have been in operation since 1954 in the Mining Department. The ball mills were installed and put in operation in 1960 in the Harding Concentrate Superphosphate Plant. The roller mills in the Mining Department are used to grind phosphate rock for outside customers. This involves short notice orders of varying grinds, grades and types of rock. The ball mills grind phosphate rock for use in V-C's own phosphoric acid and concentrated superphosphate plant. These mills grind generally two types of rock for one grind. In 1959, a 10% speed-up was made in the operating speed of the roller mills. This paper deals with the operational gains and resulting cost savings which evolved from this speed-up, and compares the present operations of the roller mills and of the ball mills in grinding phosphate rock. In January 1959 it was decided to determine the feasibility of reducing the grinding cost on the roller mills, per ton of rock ground, by speeding up the mill 10%. At this time, one mill was completely reconditioned so it would be in top condition. Extensive tests were then run both before and after the mill speed-up to determine the resultant cost savings. These tests showed that, due to this 10% speed-up, the mill capacity was increased approximately 20% and grinding power costs per ton were reduced approximately 10%. Table I shows the results of these tests. Each test run actually represents several of the same type rock and grind, averaged together. Each test run lasted approximately three hours. The validity of these runs has been confirmed by subsequent operation in grinding over 1,000,000 tons of phosphate rock. Accurate records have been kept on power consumption and repairs as well as tons produced. Two types of feed rock were used for the tests — pebble phosphate rock and flotation phosphate concentrate rock. The pebble phosphate rock size is approximately 80-90% plus 20 mesh. The flotation concentrate phosphate rock size is approximately 15% plus 20 mesh and 95% plus 100 mesh. The same two types of rock feed were used for all the test runs. Test Nos. 1, 2, and 3 were run at the original mill speed of 85.5 rpm on the vertical shaft. Test No. 1 was run with a feed of high grade flotation concentrates; the desired fineness attempted was 70% passing 200 mesh. Tests No. 2 and No. 3 were run with a feed of low grade pebble. Test No. 2 was run for a desired fineness of 70% passing 200 mesh. Test No. 3 was run for desired fineness of 50% passing 200 mesh.

Jan 1, 1963

By K. E. Merklin, F. D. DeVaney

Students of the modern blast furnace seem unanimously agreed that they are observing a major revolution in practice. Rather than changing construction and operation of the furnaces, most of the great advances now under way deal with the raw materials feed. Prepared burdens have gained increasing significance during recent years. Sizing, concentration, and agglomeration have all become bywords to greater production and reduced operating costs. The remarkable results with test furnaces in which self-fluxing sinter makes up part or all of the ore charge foretell even greater improvements. Other investigators have reported many of these advances, and it is enough to predict here that more and more operators will turn to the use of self-fluxing sinter. This change in the operator's requirements comes at a time when one major section of the iron ore regions, the Mesabi, is finding it difficult to maintain, let alone improve, the tonnage and grade of its shipping product. To meet the new demands with respect to quality and structure, the iron ore producers of this area turned first to screening, crushing, and concentrating intermediate ores. More recently, declining higher-grade reserves have led to multimillion-dollar investments in plants to treat the low-grade magnetic taconites.1,2 To yield a suitable concentrate, the Minnesota magnetic taconites must be ground to about 80 pct through 325 mesh, too fine to be charged directly to the blast furnace or to make acceptable feed to the sinter strand. The pelletizing process was developed to meet the structure requirements laid down by the blast furnace operators. Taconite producers are succeeding in their efforts just when the attention of the blast furnace man is being drawn to the great advantages of self-fluxing sinter.3 The taconite operators are ready with an alternative. Why not produce sclf-flus ing pellets? At the Hibbing laboratory of Pickands Mather & Co., where much of the early work was done that led to Erie Mining's 7.5-million ton taconite plant at Hoyt Lakes, Minn., preliminary investigations have been made. Present Taconite Operations: Details of this work will be clarified by a brief review of agglomeration practice in commercial taconite plants. After the taconite has been ground to about 80 pct through 325 mesh to liberate the magnetite grains, the ore is concentrated on magnetic separators, thickened, and filtered to about 10 pct moisture. This structure and moisture fit in closely with the conditions required to produce good balls. It has been found with taconite concentrates that at least 60 pct of the feed to the balling drum must be —325 mesh to obtain good ball compaction within reasonable time. The filter cake moisture usually permits a slight water addition at the balling drum, and further balling control is obtained by adding about 12 Ib of bentonite and 1 1/2 lb of soda ash per dry short ton of concentrate. These additives, and occasionally small quantities of fine coal, are thoroughly mixed with the filter cake ahead of balling. The swelling property of bentonite improves the green and dry strength of the balls so they can be handled without breakage until they are fired. Soda ash is added for pH control to insure maximum performance of the bentonite, and coal is added for more heat in the firing process when required. The balls, normally 3/8 to l-in. diam, are brought to a temperature of 2350" to 2500°F in the indurating furnace. Heat for this process comes from three sources: 1) fuel burned in the combustion chamber, 2) oxidation of the magnetite concentrates to hematite, and 3) coal added to the balls. Pellet strength is developed by grain growth and bridging and, if enough slag-forming constituents are present, by slag bonding. Conversion to hematite is usually close to 100 pct in the fired pellets. Batch Test Procedures: Initial studies of self-fluxing pellets were made by producing balls on a batch basis under carefully controlled laboratory conditions. The l-in. balls, uniform in size, were fired in an oxidizing atmosphere in a standard laboratory muffle furnace. The firing procedure consisted of

Jan 1, 1961

By Qingwei Wang

Heavy metal-containing wastewater from hydrometallurgical processes of lead-zinc was treated by biologics in a 200 m3/h industrial scale experiment. The result shows that Zn concentration declined from 50.28 mg/L~240.81 mg/L to 0.21 mg/L~1.98 mg/L, Pb from 1.00 mg/L~13.47 mg/L to 0.083 mg/L~0.71 mg/L, Cu from 0.24 mg/L~2.38 mg/L to 0.059 mg/L~0.40 mg/L, Cd from 2.12 mg/L~23.47 mg/L to 0.011 mg/L~0.071 mg/L and As from 0.50 mg/L~6.00 mg/L to 0.005 mg/L~0.10 mg/L. The concentrations of the above heavy metals in the treated water were lower than that in Integrated Wastewater Discharge Standard (GB8978~1996) in China. Furthermore, the treated wastewater was recycled and reused in smelter plant. Zinc content in hydrolytic sludge reached up to 34.04%, which indicates that this sludge can be used to recover valuable metals.

Jan 1, 2010

By L. A. Cramer

Platinum metal was first utilized by the Inca civilization in South America where it was found in alluvial deposits along the San Juan and Atrato rivers that lie between the Andean mountain range in Colombia and the Pacific Ocean. The nuggets found in these rivers average between 77% and 86%platinum, 7%?8% iron and small amounts of copper, gold and other Platinum Group Metals (PGMs). The Incas often alloyed platinum with copper to increase its workability and fabricated various ornaments from it. With the arrival of the Spanish in the early 1700s, platinum was labelled a nuisance for the Spanish in their quest for alluvial gold because it concentrated with the gold and had to be removed by either extended amalgamation or by hand sorting. It was at this time that platinum received the root of its current name for the Spanish referred to it as platina del Pinto, platina being the somewhat derogatory diminutive of plata (silver). Pinto was a river in Colombia where the Spanish first distinguished it. The English eventually adopted ?platina? and it is only in recent times that the Latin name platinum has been accepted as the metal?s common English name. From its first introduction to European scientists in 1740, platinum was unappreciated because of its brittle nature and difficulty of working and as a result had little commercial value. Its ability to be dissolved in gold did lead to some counterfeiting of gold bullion and even to the debasement of gold currency.

Jan 1, 2000

By Fabian Kirsch

Cyclic behaviour of axially loaded piles has been an issue since studies revealed that the capability to transfer skin friction into the ground is highly depending on the level and the number of load cycles the pile is subjected. It has been shown that degradation of skin friction due to axial cyclic loading leads to accumulating displacements and a severe reduction in pile capacity. In Germany the DIN 1054 as well as the Recommendations of the Working Group 2.1 ?Piles? of the German Geotechnical Society (EA-Pfähle) require the consideration of cyclic loading for the assessment of both the displacements and the capacity. Up to now no broadly accepted design method exists, which enables the engineer to give a reliable judgement on the pile capacity under axial cyclic loading and the displacements respectively. Instead numerous design charts make it possible to distinguish between stable and instable loading levels in order to ensure a design on the safe side. A newly developed design approach based on engineering models is presented which delivers a calculation algorithm for the remaining axial pile capacity after a defined number of load cycles to allow for the analysis of stability of the foundation. Furthermore the computation of accumulating displacements is made possible. The results are verified with pile load tests undertaken at Dunkirk, France and showed that the axial capacity after cyclic loading calculated with the newly developed design approach predicts actual capacity with an accuracy of ±20%.

Jan 1, 2011

The sedimentary basins of this region developed in the collision zone of thePacific Ocean plate and the continent. Stresses were transmitted thorough wedge-like lithospheric blocks, such as the Okotsk Sea, East Kamchatka, Aleutian and other smaller blocks. Compression and inversion were more active at the boundaries: ***INSERT EQ*** and ***INSERT EQ*** in places. Tension, rifting and the increase in the rate of sedimentation were marked in P1-2, P3, N12, and also in N2 û Q in places. Structural plan changes of the sedimentary basins and possible oil, gas and fluid migration occurred more actively at the boundaries: ***INSERT EQ***.

Jan 1, 1995

By C Ross

This paper uses the results of research trials by Landcare Research and Forest Research Institute and assessments of reclamation plantings at mines and other drastically disturbed sites to describe methods of enhancing restoration of native forests. The characteristics of mine and forest environments are compared to highlight the difficulties facing restoration and techniques which result in successful restoration are described. Selection of the most appropriate technique is illustrated by describing restoration of lowland beech-podocarp forest at a West Coast coal mine. The key to successful reclamation is to understand the factors controlling and limiting plant establishment and growth, then to design rooting media, surfaces and management regimes to create a favourable environment. This paper concludes with a summary of the planning and technical factors which are critical for successful restoration of native forest.

Jan 1, 1995