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By Earl M. Frizzell, Joseph A. Dos Santos

"Conventional belt conveyors offer an economical method for transporting materials at inclines approaching 18 degrees. Sandwich belt conveyors may be used to transport materials at high angles, to 90 degrees, while maintaining the most positive features of conventional conveyors, that is, (1) proven conventional conveyor components. (2) high-speed operation, (3) high-single lifts, (4) continuous cleaning of the smooth surf aced belts, and (5) easy, quick repair by hot or cold vulcanizing. The present discussion deals with the evolution and development of sandwich belt conveyors to satisfy the operational requirements of the open-pit mining and materials-handling industries.A look at the limitation s of conventional shallow angle conveyors reveals the conveying angle at which a cover belt is needed. The relationship between conveying angle and hugging pressure by the cover belt is developed mathematically. A study of past and current developments reveals that the Loop Belt and Beltavator are significant advances in the state-of -the- art. The vertical radius of curvature constraints are discussed in detail and the constraint equations are developed. Criteria for future development is established as a basis for evolutionary developments in sandwich belt high-angle conveyors. The use of intermediate drives in sandwich belt high-angle conveyors is investigated.IntroductionIn connection with U.S. Bureau of Mines funded–study relating to continuous haulage from deep open-pit mines, the authors had occasion to study extensively, the use of high-angle conveyors as a means of hauling mine products continuously from the pit while maintaining optimum mine slopes to minimize total excavation. The many possible high-angle conveying methods investigated included skip hoists, bucket ladders, pocket belts, fin belts, cleated belts, sandwich belts, pipe belts, screw conveyors, slurry pipe lines, and others. The sandwich belt high-angle conveyor appeared to be the most operationally-appropriate and economical solution for the mining industry. The present paper presents the theory, historical development, state-of-the-art, and evolutionary developments in sandwich belt high-angle conveyors."

Jan 1, 1983

By Frank E. Chase

Retreat mining is one of the most hazardous operations conducted underground. One of the keys to miner safety and an efficient recovery of the reserves is to design adequately sized production pillars to prevent pillar squeezes, excessive pillar spalling, severe floor heave, roof falls, and pillar bumps. Currently, few pillar design approaches are available which consider the abutment pressures generated by the gob. The Bureau of Mines is developing a method called Analysis of Retreat Mining Pillar Stability (ARMPS) to aid in the design of retreating sections. Back analysis of 42 case studies indicates that ARMPS can be helpful in predicting stability during retreat mining operations.

Jan 1, 1993

By I. H. Clark

A variety of backfills is presently being placed in different gold mines. The backfills are pumped as slurries to stopes where they are discharged into some form of retaining paddock. The effectiveness of fill as a support medium will depend on a number of factors. These can largely include (i) inherent properties of the material, e.g. particle size, shape, distribution and com- position, and (ii) conditions of placement. The terminology required for describing backfills is introduced, and the various laboratory-measured properties are presented for a variety of fills. Finally, the properties of the fills as they are placed underground are discussed.

Jan 1, 1988

By P. Vlasak, Z. Chara

Since the flow behaviour of slurry containing both coarse and very fine particles has not been sufficiently clarified, the paper focuses on the experimental study of the particle size distribution, time and intensity of shearing and the effect of chemical agents on the hydraulic gradient versus the average velocity relationship of fluidic ash slurries in the laminar and turbulent regimes. The recirculation pipe loop with smooth stainless steel pipes of inner diameters 26.8 and 36 mm was used for measuring the ash-water mixtures flow parameters. The dense fluidic ash-water mixtures behave as time dependent non-Newtonian slurries. Shearing, the addition of bottom ash or chemical agents, favourably affects the flow behaviour of dense fluidic ash-water slurries. The hydraulic gradient in the laminar region significantly decreases; also the laminar/turbulent transition region is reached for lower flow velocity.

Jan 1, 2007

By J. Moore Samuel

THE prospective change in ore receipts at the Copper Queen reduction works necessitated a careful study of conditions to determine the most economical method of smelting. The first step, calculating the known, probable, and possible results, showed the economy of a unimetallurgical flow sheet f which the three pyrometallurgical stages would be roasting, reverberatory smelting, and converting. The past and future ore receipts (not including converter flux) are as follows: AVERAGE MONTHLY FUTURE MONTHLY RECEIPTS, 1918 RECEIPTS TONS PER PER CENT. OF TONS PER PER CENT. OF MONTH TOTAL RECEIPTS MONTH TOTAL RECEIPTS Ore 84,553 82.3 43,000 44.0 Gravity concentrates 16,375 15.9 25,500 26.0 Flotation concentrates 1,883 1.8 29,400 30.0 Total 102,811 100.0 97,900 100.0 Of the 102,811 tons per month received in 1918, only 28,945, or 28 per cent., was treated in the roasters and an average f 14.1 roasters out of 20 were in operation; the remaining ore was treated in the blast furnaces. The change to "all reverb." practice, therefore, called for considerable additional roaster capacity, which could be obtained either by building additional units, or by remodelling the present units so as to give increased unit capacity. In view of the progress in roasting practice since the construction of the Copper Queen roaster plant and the heavy investment in that roaster plant, the second plan seemed to be the more desirable. The Copper Queen roaster plant consisted of twenty six-hearth, McDougall roasters, 18 ft. in diameter, equipped with cast-iron arms that originally were water cooled but now are air cooled. They have no top dry hearth and the feed is discharged through the top arch, near its periphery, by plunger feeders from two bins at opposite sides of the roaster.

Jan 7, 1921

By L. C. de Kooker, K. Esmaeili, I. Musonda, M. Ferentinou

A numerical investigation of the effect of pore pressure regime on the safety factor and the critical failure mechanism is presented for fly ash storage facility. Pore pressures’ measurements from standpipe piezometers and pore pressure estimated from seepage analysis are used to compare the factor of safety for a fly ash slope. This was applied for considering static and seismic scenarios. A probabilistic approach was applied to account for the uncertainties resulting from the limited data available and support a qualitative risk assessment evaluation. Slope stability analysis is conducted in two and three dimensions, adopting the limit equilibrium analysis approach, and also a finite element seepage analysis, to assess the stability of the slope. The two-dimensional cross-sections were extruded to three-dimensional models to estimate the factor of safety and associated shear failure. The results from the performed analysis suggest an increase in safety factor values of 5%.

Mar 27, 2024

By W. A. Gow, J. A. Vezina

"A new process for treating scheelite for the production of tungstic tri-oxide is described. The novelty of the process resides in that an ion-ex-changer is used to convert sodium tungstate to either ammonium tung-state or tungstic acid.The process consists of a conventional hydrochloric acid leach of the scheelite concentrate, and the dissolution of the impure tungstic acid thus produced with sodium hydroxide at pH 8. The sodium tungstate solution is clarified and passed through a column of Dowex 50Wx8 ion-ex-change resin to exchange the sodium ions for either hydrogen or ammonium ions, both being suitable. The tungsten-bearing column effluent is evaporated to crystallize the tungsten as ammonium tungstate; the crystals are washed with nitric acid and water and decomposed by heat to pro-duce tungstic trioxide of high purity.Ninety-six per cent of the tungsten in the concentrate may be re-covered as tungs tic trioxide (W03), and the reagents consumed are 0.7 g HCl, 0.41 g NaOH and either 1.73 g NH.Cl or 225 g HCl per gram of tungsten."

Jan 1, 1966

"Mining with blast holes drilled upwards may be done to recover sill pillars, or for economic reasons when the distance between the levels is relatively large. If the rock is sufficiently competent to accommodate large stope size, the distance between sublevels could be increased to as much as drilling accuracy allows. The same drill drift can be used for the drilling of both down holes and up holes. In the case of down holes, the initial opening of the stope (drop raise) is not as much a problem as for the inverse drop raise for mining with up holes. This is because, with a drop raise, a lot of space is available at the toe for the blasted rock to expand. In addition, the raise can be blasted in many shots, thus allowing smaller blasts each time. Inverse drop raises are generally much more difficult to blast because of the limited space for muck flow, and must, therefore, be taken down in a single blast, requiring special considerations in order to ensure a success. A case study of a 29 m (95 ft) inverse drop raise successfully blasted in a single shot at Heath Steele Mines is described. The paper covers technical considerations such as drilling accuracy, hole deviation survey, delay extension and protection, initiation methods and initiation sequence. Finally, a novel method for creating the initial opening called ’slot drilling’ is presented. This method, inspired from LKAB in Sweden, consists of drilling a row of interconnected holes using a special guide mounted on a regular production drill hammer.With a case study of blasting an inverse drop raise of 29 m (95 ft) in one blast, this paper further discusses some technical issues that are critical for the development of inverse drop raises. A novel method for creating the initial opening with a slot of interconnected holes is also presented with the experience acquired at different mine sites across the Noranda group."

Jan 1, 2000

By Packer M. L

Residual load development in driven precast piles has been widely researched in recent decades, however, cast-in-situ piles are often falsely assumed to be unaffected by this mechanism. In contrast to driven precast piles, driven cast-in-situ (DCIS) piles require adequate time prior to loading for the concrete hydration and curing processes to occur. The concrete-soil interaction is complex with several processes influencing the strains within the concrete pile after installation. DCIS piles are of particular interest given that residual loads can be developed during both the driving and casting phases. This paper present four instrumented end-bearing DCIS piles in which temperature and strains were monitored from installation, with the objective of identifying the processes contributing to the measured strains prior to testing. Strain gauges at all depths recorded a net compression and compressive strains continued to increase up to the commencement of load testing. Residual strains in the test piles were interpreted to be influenced by temperature (concrete and ambient soil and air temperature), consolidation of soils due to dissipation of excess pore water pressure, and swelling from absorption of groundwater. Several residual load interpretive methods are evaluated with the Siegel and McGillivray (2009) approach being consistent with the skin friction full mobilisation theory of the Unified Design method. The full-scale field tests presented herein provide a unique opportunity to examine DCIS pile behaviour, particularly given the lack of presently available published studies on these pile types.

May 1, 2022

By James A. Robertson

"The Pronto Mine located at the east end of Lake Lauzon, Long Township, some 12 miles east of Blind River is of historic note being the discovery locality and the first producing mine of the Elliot Lake uranium camp. Radioactive bedrock was discovered in 1949, a commercial deposit was verified in 1953 and production commenced in 1955. Production continued to April 1960: total tons milled 2,264,404 from which 4,643,835 pounds U,O, were recovered at a recovered grade of 2.05 pounds U3O8 per short ton.The orebody is in a bed of pyritic quartz pebble conglomerate at the base of the Matinenda Formation of the Elliot Lake Group of the Huronian Supergroup. The rocks strike east and dip 15 degrees to 20 degrees south and unconformably overlie Early Precambrian granite, the upper surface of which is marked by a regolith. Several diabase dykes cut the orebody which is also affected by albite, chlorite, and carbonate-rich alterations. To the south the orebody is cut by a fault. The orebody is 6to 10 feet but locally 10 to 15 feet thick, has a strike length of 3,500 feet and extends down dip for 3,600 feet to a depth of 1,000feet. Some 300,000 tons of ore grade mineralization remain in the deeper portion of the mine.IntroductionThe Pronto Mine, (Photo I) at the east end of Lake Lauzon in Long Township some 12 miles east of Blind River, is of particular interest as being the discovery site and the first producing mine of the Elliot Lake camp.Production at the Pronto Mine commenced in 1955 and continued until 1960 when the mill was converted to treat copper ore from the nearby Pater Mine at Spragge. A description of both mines, based on material provided by the mine staff, was given in the report on ""Geology of the Spragge Area"" (Robertson, 1970).As little study has been done on Pronto geology or petrology since mine closure, the earlier description of exploration and geology is essentially repeated, but shortened, here- the main change being that the stratigraphic terminology has been amended as recommended by the Federal-Provincial Committee on Huronian Stratigraphy (Robertson et al., 1969)."

Jan 1, 1986

By D. J. Jr. McAdam

PART I.-OUTLINE OF INVESTIGATION, DESCRIPTION OF METHODS AND MATERIAL Previous Investigation of the Influence of Stress on Corrosion IN 1917 Haigh1 presented evidence that under simultaneous corrosion and cyclic stress metals may fail at lower stresses than if the corrosion is prior to the cyclic stress. In 1926 the author, while investigating the cooling effect of water on the fatigue resistance of metals, found that under these conditions steel specimens fail at stresses far below the endurance limit.2 (By "endurance limit" is here meant the fatigue limit obtained with specimens as free as possible from stress concentration and corrosive influences.) This led to an investigation (at the Naval Engineering Experiment Station, Annapolis, Md.) of the influence of cyclic stress on corrosion. This investigation maybe divided into three parts. The first part was a general survey of the conditions necessary to cause complete failure under corrosion and cyclic stress of relatively high frequency. The entire process by which metals fail under these conditions was-called "corrosion-fatigue." "Corrosion-fatigue limits" for a large number of metals were determined and the relationships of these limits to chemical composition, heat treatment and physical properties were established. The conclusion was reached that corrosion-fatigue is due to the accelerating influence of cyclic stress on corrosion. Results of this part of the investigation were presented in five papers.2-6 The second part of the investigation was a preliminary study of the influence of cyclic stress on corrosion. Each experiment was conducted in two stages. The first was a corrosion stage in which the specimen was corroded under cyclic stress; the second was a fatigue stage, in which the previously corroded specimen was tested (without further corrosion) to determine its fatigue limit. The difference between this fatigue limit and the endurance limit was used as a criterion of the "damage" due to

Jan 1, 1931

By Filipe Sathler

O Alto Forno 3 da ArcelorMittal Tubarão (Volume interno: 3617 m3) esteve em operação do dia 21 de Julho de 2007 a 10 de Novembro de 2012 durante a sua primeira fase da primeira campanha. Após os devidos reparos e planejamento técnico-operacional, o procedimento de Blow in da segunda fase desta campanha foi realizado no dia 6 de Julho de 2014, apresentando também um processo bem sucedido de evolução de produção (ramp up) e resultados importantes com a redução do consumo de coque (coque rate) ao longo dos primeiros meses de operação.

Aug 17, 2017

By Wils L. Cooley

Two major grounding topics are covered: mine ground systems and ground-check monitors. Performance characteristics of ground-check monitors for high-voltage surface mine distribution cables and low-voltage underground mine trailing cables are evaluated. Over 20 commercially available monitors are studied to determine their utility. Several monitoring concepts are incorporated into an unusual monitor design, the prototype of which is described in detail. A comparative study of the effectiveness of several alternate methods of grounding surface mine equipment supports the system currently proposed by MSHA. An instrument design to measure earth voltage gradients is described, as well as an instrument design for measuring the effectiveness of dc haulage current return systems. An extreme analysis is made of techniques for measuring earth resistivity and resistance of safety ground beds. Electrolytic corrosion studies of earth electrodes are reported extensively also.

Jan 1, 1979

By Erika Moonin, Jim Nickerson, Claudio Cimiotti, Roberto Bono

"In response to the severe drought on the Colorado River basin and in order to preserve existing water capacity to the Las Vegas Valley, the Southern Nevada Water Authority contracted a new deep-water intake (Intake No. 3) located in Lake Mead. The current project (Contract No. 070F-01-C1) includes 5 km (3 miles) of tunnel with very difficult geological conditions and a 200- m (650-ft) deep shaft and marine works. This paper will present the current status of the design-build contract being constructed by Vegas Tunnel Constructors and the challenges and problems faced and the innovative solutions developed to handle the difficult conditions encountered during tunneling with up to 15 bars of pressure.ProjectbackgroundThe Las Vegas Valley receives the majority of its water from Lake Mead, which is located approximately 30 km (20 miles) east of the Las Vegas, NV metropolitan area. Due to the severe drought that has hit the southwest region of the United States, the water level in Lake Mead has receded from its high water level of 372 m (1,221 ft) above mean sea level (amsl) to a low of 330 m (1,076 ft) amsl in June 2015. Should the lake level continue to fall, at elevation 320 m (1,050 ft) amsl, the current water intakes will not be able to draw enough water to meet the needs of the local area. The risk of losing water capacity has prompted the construction of a new deeper intake. Not only is the new Intake No. 3 situated at a lower elevation in the lake at 262 m (860 ft) amsl, the new Intake No. 3 is also located to draw better quality (Fig. 1).Project challengesThe project has had several difficult challenges to overcome:• The work consisted of a 1.2 kt (1,300 st) structure constructed on a barge and lowered 100 m (330 ft) into Lake Mead anchored with 9,200 m3 (12,200 cu yd) of tremie concrete.• The drill-and-blast starter tunnel was impacted by three inflow events, resulting in a one-year project delay.• One of the main challenges was selecting the tunnel boring machine (TBM) and excavating the tunnel with expected hydraulic head pressure of as much as 17 bar (16 bar was the highest pressure recorded during the drive), crossing faulted areas, low rock cover and the risk of tunnel instability due to direct connection with the lake.• Extensive pre-excavation grouting with a constant head pressure ranging from 12 bar to 15 bar.• Structural repair of the cutterhead in difficult hydrogeological conditions [water inflow of approx. 800 m3/hr (4,000 gpm)].• Complete repair of the cascade sealing system for the main bearing and two replacements of the pinion bearings.• The TBM approach to the intake structure."

Jan 1, 2015

By G. L. Oldright

"The various districts in the United States where complex lead-zinc ores occur are too well known to require description, there being hardly a mining State with any large amount of ores of either metal that is free from this problem. This paper briefly presents developments in the volatilization process, as applied to the treatment of such ores, since the data in Bureau of Mines Bulletin 211** were published. In that bulletin, among the many types of ore treated, examples are given of the treatment of lead-zinc ores by the chloride volatilization process for the separation of lead. An outline of the general process, with views of the rotary kiln, ore feeder, firing system (these three features resemble closely those of cement-kiln practice), dust chambers, and a Cottrell precipitator to remove the fume from the gas, are also shown in that bulletin, to which the reader is referred for the results obtained, and for practical details of operation. The small initial cost of a volatilization plant, and its adaptability for use in isolated districts were emphasized.In general, it was found that when an intensely oxidizing atmosphere was maintained in the furnace, a good lead-zinc separation could be made. Therefore, a heavy sulphide ore would have to be preroasted, and enough air would have to be admitted to obtain complete combustion of whatever fuel might be used.The work done by private companies and the bureau since the data for Bulletin 211 were obtained, indicates that the ores most immediately amenable to treatment by volatilization are of two types. The first is an oxidzed lead. silrer ore whose variable lime content makes it impractical to chloridize the ore for acidulated brine leaching, and whose base metal content makes treatment by cyanidattion unfeasible, The second type is the lead-zinc ore mentioned above, and is the one on which data are available for publication at present. This article Chows what ores of this latter type are amenable to treatment by volatilization, by giving briefly some of the more important physical and chemical properties an ore should preferably possess. Several examples are given from both laboratory experimental work, and from practice in the field, showing various phases of the volatilization process."

Nov 1, 1923

By R. P. Bhagat

The paper reports results on sinterability of iron ores from the mines of Tata Steel Limited. The objective was to improve tumbler index (TI) of sinter and its reduction parameters, reduction degradation index (RDI) and reducibility (RI) while maintaining strand productivity same or to improve. Pot sintering experiments were carried out with Joda and Noamundi mines having LOI 3.8% and 2.9% respectively. Sinter basicity (CaO/SiO2) was varied in the 2.1 to 2.5 range and its MgO content was kept in the 1.4% to 1.8% range according to the blast furnace requirement. The ignition condition and vacuum was maintained according to the plant practice. The effect of variables namely, granulometric composition of ores vis-à-vis that of sinter mix, bed height, solid fuel, moisture content of green mix on the sintering indices was studied. The productivity with Noamundi ore (26 %+6.3 mm) was lower than that with Joda ore (45% +6.3 mm) at 4% coke in the mix and sinter basicity (CaO/SiO2) 2.1. However, with 4% coke (ash 22%) the RDI value was significantly high in case of Noamundi Ore, which decreased significantly to 29% with 5% coke. The size parameter of ore fines and its LOI content do influence the sintering speed and the sinter properties significantly. At 2.1 basicity ratio, the sinter with the Joda ore having coarsergranulometry had highest TI (=70.2%). The RDI (27%) in this case was inferior to that (23.2%) when Joda ores with finer size was sintered. Better RDI with finer sized ore could be attributed to better assimilation of CaO/ MgO with more balanced and uniformly distributed hematite minerals. The size parameter has been inter-influenced by the coke requirement. With the lower mean sized Joda ore the sinter RDI decreased significantly to 23% and the productivity improved to 42.9 T/(sq m*d) when the coke in the mix increased from 4% to 5%. An attempt to reduce the coke in the mix from 5% to 4% with respect to both the ores was not successful as it resulted higher RDI and lower strand productivity. The present study has revealed that sinter RDI improves (decreases) with decrease in the sintering time, while such correlation was absent between TI of sinter and speed of sintering. Corollary to this observation, one can take advantage of other stronger parameters to increase the sintering speed (and hence sinter productivity) without affecting the sinter strength. In the present study higher FeO sinters has displayed excellent RDI, however cold strength of sinter was not correlated with its FeO content. The TI and RI parameters of sinter improves marginally with increase in its MgO content while RDI got worsened. Keywords: LOI content, iron ore, reduction properties, sinter, noamundi, joda

Sep 1, 2012

By Jens Classen, Martino Gatti

"GENERAL PROJECT DATAThe Sparvo Tunnel is part of the “Variante di Valico” project, a new 65km (40mi) highway link between Bologna and Florence, designed to renew the existing A1 route which connects Bologna and Florence. It is composed of 2 x 2,5km (1,55mi) twin tunnels excavated inside a complex geological context, characterized by the presence of a series of surface landslides and methane gas.The inclination is 2,6 % and the minimum radius is 1.400m (4.600ft). The two tunnel tubes were excavated by a 15,6m (51ft) diameter EPB-TBM from summer 2011 to summer 2013. The internal diameter of the concrete lining is 13,6m (44,5ft), made up of precast concrete segments 2m (6,5ft) long and 70cm (28”) thick, being the largest and heaviest segments ever produced. Excavation went smooth apart from initial problems concerning clogging problems and a main bearing damage of the TBM, which are discussed in a further paper during the congress. No structural damage was observed until the end of the TBM drives.SITE OBSERVATIONS DURING THE POST-EXCAVATION PHASEThe first evidence of the stress state of the segmental lining appeared on November 20, 2013 inside the northern tube (1st drive). Ring No. 884 showed obvious signs of overstress, resulting in spalling and crushing of the concrete, and the subsequent bending of the reinforcement, as visible in figure 1.At this time the tunnel invert was already partially backfilled, and construction traffic was running through the tunnel. First safety measures included closing of the tunnel, and initial measurements of the displacements."

Jan 1, 2016

By F. A. Rutledge

Mercury deposits were discovered in 1941 in the sedimentary rocks of the Cinnabar Crook area, near the head of the Holitna River drainage in southwestern Alaska, by Russell Schaefer and Harvey Wincholl of Sloitmut, Alaska. High-grade float was found in both Canary Gulch and Cinnabar Gulch, and indications of other deposits wore found in Aldor and Broken Shovel Gulches. Schaefer and Winchell's exploration in 1942 and 1943 revealed placer deposits of cinnabar along Canary Gulch. During prospecting on Canary Gulch, 3,600 pounds of ore was recovered from the detrital material on the Lucky Day lode at the head of the gulch. When retorted, this ore was reported by the owners to have produced 26 flacks of quicksilver. During the summer of 1943, Wallace M. Cady and Charles A. Hickcox, of the U. S. Geological Survey, spent 4 weeks mapping and investigating the area. Bruce I. Thomas and Harold C. Pierce, mining engineers of the Bureau of Mines, examined and sampled several deposits in September 1943. Following their preliminary examination, in view of the exceptionally high-grade cinnabar present, additional investigation by the Bureau of Mines was recommended. A program of trenching, test pitting, and sampling to delimit the deposits was completed during the summer of 1947 under the direction of the author.

Jan 1, 1950

By G. E. Johnson

A bench-scale study was made of removing hydrogen sulfide from synthesis gas at elevated pressure by iron oxide impregnated on wood chips. The gas contained 0.02 to 0.06 percent oxygen. The removal of hydrogen sulfide by iron oxide impregnated on wood chips is more efficient at elevated pressure. At constant space velocity, longer periods of operation at 400 p.s.i.g. are possible before the exit gas reaches a given maximum concentration of hydrogen sulfide, compared with 200 p.s.i.g. or atmospheric pressure. In tests at room temperature, 60° C., and 100° C., the most effective removal of hydrogen sulfide was obtained at room temperature and the least effective at 100° C. Dry impregnated wood chips removed only about half as much hydrogen sulfide as chips wetted to a water content of 50 percent.

Jan 1, 1962

By Donald E. Redmon

This Bureau of Mines report reviews injury experience in the nonmetallic mineral industries (except stone and coal) for 1966-67. For each of the years, 23 tables summarize disabling work injuries by department, source of injury, extent of disability, part of body injured, and nature of injury. Correlative information on employment, worktime, and operating activity is also presented.

Jan 1, 1973